TN 295 



No. 9135 




ureau of Mines Information Circular/1987 



Surface Mine Blasting 

Proceedings: Bureau of Mines Technology Transfer 
Seminar, Chicago, IL, April 15, 1987 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 



Information Circular 9135 



Surface Mine Blasting 

Proceedings: Bureau of Mines Technology Transfer 
Seminar, Chicago, IL, April 15, 1987 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 
Donald Paul Hodel, Secretary 

BUREAU OF MINES 
Robert C. Horton, Director 




Library of Congress Cataloging in Publication Data: 



%<& 






f 



.o- 



Bureau of Mines Technology Transfer Seminar (1987 : 
Chicago, m.) 

Surface mine blasting. 

(Information circular ; 91 35) 

Includes bibliographies. 

Supt. of Docs, no.: I 28.27:9135. 

1. Blasting— Congresses. 2. Mining engineering— Congresses I. United States. Bureau 
of Mines. II. Title. III. Series: Information circular (United States. Bureau of Mines); 
9135. 

TN295.U4 [TN279] 622 s [622'.31] 87-600079 



PREFACE 

This Information Circular (IC) summarizes recent Bureau of Mines re- 
search to enhance the efficient and safe use of explosives in mining. 
Many of the papers in this publication were presented at a Bureau of 
Mines Technology Transfer Seminar on Surface Mine Blasting on April 15, 
1987, in Chicago, IL. The papers in this IC represent only a small por- 
tion of the Bureau's overall research effort to improve current raining 
technology and the health and safety of mine workers. Information about 
other research programs or technology transfer activities aimed at in- 
troducing these research results to the minerals industry may be ob- 
tained by contacting the Bureau of Mines, Branch of Technology Transfer, 
2401 E Street NW, Washington, DC 20241. 



CONTENTS 



iii 



Page 



Preface i 

Abstract . . 1 

Bureau of Mines Surface Mine Blasting Research, by Dennis V. D'Andrea 2 

Reducing Accidents Through Improved Blasting Safety, by Larry R. Fletcher and 

Dennis V. D'Andrea 6 

Blaster's Training Manual for Metal and Nonmetal Miners, By Michael A. Peltier, 

Larry R. Fletcher, and Richard A. Dick 19 

Delayed Blasting Tests To Reduce Rockfall Hazards, by Virgil J. Stachura and 

Larry R. Fletcher 25 

Effects of Blast Vibration on Construction Material Cracking in Residential 

Structures, by Mark S. Stagg and David E. Siskind 32 

Blast Vibration Measurements Near Structures, by David E. Siskind and 

Mark S. Stagg 46 

Initiation Timing Influence on Ground Vibration and Airblast, by John W. Kopp.. 51 
Vibrations From Blasting Over Abandoned Underground Mines, by David E. Siskind 

and Virgil J. Stachura 60 

Computer Modeling of Rock Motion, by Stephen A. Rholl 73 

Influence of Blast Delay Time on Rock Fragmentation: One-Tenth-Scale Tests, 

by Mark S. Stagg and Michael J. Nutting 79 

Blasting Effects on Appalachian Water Wells, by David E. Siskind and 

John W. Kopp 96 

Fiber Optic Probe to Measure Downhole Detonation Velocities of Explosive 

Columns , by David L. Schulz 103 

Stemming Ejection and Burden Movements of Small Borehole Blasts, by 

John W. Kopp , 106 





UNIT OF MEASURE ABBREVIATIONS 


USED IN 


THESE 


PAPERS 


atm 


atmosphere 


(in/s 


)/in 


inch per second 
per inch 


°C 


degree Celsius 


K 




kelvin 


dB 


decibel 


kHz 




kilohertz 


°F 


degree Fahrenheit 


lb 




pound 


ft 


foot 












lbf/in 2 


pound (force) 


ft 2 


square foot 






per square inch 


ft/lb ,/2 


foot per square root 
pound (scaled distance) 


m 
mg/L 




meter 

milligram per liter 


ft/lb 1/3 


foot per cube root 










pound (scaled distance) 


mi/h 




mile per hour 


ft/s 


foot per second 


rain 




minute 


G 


gravity (32.2 ft/s 2 ) 


yin/in 


microinch per inch 


g 


gram 


mm 




millimeter 


gal/d 


gallon per day 


ms 




millisecond 


gal/min 


gallon per minute 


ms/ft 




millisecond per 
foot 


(gal/min)/ft 


gallon per minute per 










foot (specific well 


ys 




microsecond 




capacity) 


mV 




millivolt 


g/cm 3 


gram per cubic centimeter 


ns 




nanosecond 


h 


hour 


pet 




percent 


Hz 


hertz 


s 




second 


in 


inch 


W 




watt 


in 3 


cubic inch 


yr 




year 


in/s 


inch per second 








in 3 /s 




cubic inch per second 









SURFACE MINE BLASTING 



Proceedings: Bureau of Mines Technology Transfer Seminar, 

Chicago, IL, April 15, 1987 



Compiled by Staff, Bureau of Mines 



ABSTRACT 

The Bureau of Mines has sponsored a comprehensive research program to 
enhance the safe, effective, and efficient use of blasting technology by 
the raining industry. Recent research results of the surface mine blast- 
ing program were presented at a seminar on April 15, 1987, in Chicago, 
IL. Many of the topics discussed at the seminar are presented in this 
proceedings. The new research described in these papers includes com- 
puter monitoring of rock motion, the influence of blast delay times on 
rock fragmentation, blasting effects on Appalachian water wells, blast 
vibration measurement near structures, and the reduction of accidents 
through improved blasting safety. 



BUREAU OF MINES SURFACE MINE BLASTING RESEARCH 
By Dennis V. D'Andrea 1 



ABSTRACT 



The Bureau of Mines Twin Cities Re- 
search Center has a comprehensive re- 
search program on the efficient and safe 
application of explosives in mining. Re- 
searchers combine an understanding of the 
basic principles of dynamic rock fragmen- 
tation with new blast design technology 
and recent developments in both methods 



and equipment, for potential improvements 
in blasting practices. This paper out- 
lines surface mine blasting research com- 
pleted since the Bureau's last Technology 
Transfer seminar on blasting in December 
1980. Three programmatic areas — produc- 
tivity technology, blasting vibrations, 
and blasting safety — are reviewed. 



HISTORY OF RESEARCH 



Blasting research has been conducted 
at the Twin Cities Research Center (TCRC) 
since the center opened in 1959. Re- 
search during the 1960's and early 1970's 
established TCRC as the leading Bureau of 
Mines center in the area of blasting for 
improved fragmentation and increased pro- 
ductivity. During the period from fiscal 
year 1974 through fiscal year 1979, pro- 
ductivity research was on blasting to 
prepare ore bodies for in situ leaching. 
The major effort at TCRC from fiscal year 
1975 through fiscal year 1983 was in the 
area of environmental effects of blasting 
(ground vibrations and airblast). Re- 
search since fiscal year 1983 has been 
mostly on blasting fundamentals for 
improvements in productivity. Blasting 



safety research began at TCRC as one con- 
tract project in fiscal year 1978 and 
grew to involve four in-house projects 
during the years 1984 to 1986. 

The heavily field-oriented blasting re- 
search program at TCRC has included 45 
in-house and contract project efforts 
since 1975, resulting in 122 publications 
and numerous presentations at profes- 
sional meetings. Report of Investiga- 
tions (RI) 8507, on structural response 
and damage from blasting vibrations, won 
the 1981 Applied Research Award from the 
U.S. National Committee for Rock Mechan- 
ics. TCRC personnel have responded to 
over 650 requests for technical assis- 
tance and advice on blasting since 1981. 



CURRENT RESEARCH PROGRAM 



Mining Technology 



Major research efforts at TCRC on im- 
proved productivity and blasting vibra- 
tions in surface mines are listed in ta- 
ble 1. Included are projects intended to 
improve mining productivity and to pro- 
vide information on good blasting prac- 
tices. The projects that started in the 
late 1970's addressed environmental is- 
sues, with indirect implications for 
mining costs and productivity. More re- 
cent long-range, high-risk research is 

' Research supervisor, Twin Cities Re- 
search Center, Bureau of Mines, Minne- 
apolis, MN. 



examining the fundamentals of blasting 
and blast-produced rock fragmentation. 
The fundamental research includes a study 
of high-precision delay initiators to im- 
prove fragmentation. 

The Bureau is frequently asked to as- 
sist other Federal agencies, such as the 
Office of Surface Mining, the Bureau of 
Reclamation, and the Mine Safety and 
Health Administration, on environmental 
and safety issues associated with blast- 
ing practices. These assistance efforts 
usually do not involve pure research, but 
rather providing technical Information 
and, on occasion, measurement, analysis, 
and advice. 



TABLE 1. - Blasting research at TCRC on improved productivity 
and blasting vibrations 



Research area 


Fiscal 


Key researchers 


Significant 




years 




publications 1 


Blasting effects on 


1978-80 


P. R. Berger and 


D. Robertson (1). 


Appalachian water wells. 




Associates. 2 




Contour mine blasting 
noise and vibrations. 


1978-81 




Stachura, RI 8892 






(2). 


Standards efforts on 


1980-81 


D. E. Siskind, Acousti- 


ANSI Standard 


vibrations, ANSI and 




cal Society of America. 


S3. 29-1983 (3). 


ISO. 






Other ANSI and ISO 
Standards. 


Fatigue from repeated 


1979-83 


M. S. Stagg, National 


Stagg, RI 8896 (4). 


blasting. 




Bureau of Standards. 2 


Siskind (5). 
Woodward (6). 


Vibration measurement 


1983 




Siskind, RI 8969 CO. 


near buildings. 








Blast designs to control 


1979-83 




Kopp, RI 9026 (8). 


vibrations. 








Blasting fragmentation 


1984- 


M. S. Stagg, Sandia 


Nutting (9). 


fundamentals. 


(ongoing) 


Laboratories, Univer- 
sity of Maryland. 2 




Low-frequency vibrations. 


1985- 


D. E. Siskind, 


In preparation. 




(ongoing) 


V. J. Stachura. 





ANSI American National Standards Institute. 
ISO International Standards Organization. 
Only the senior author is listed here. Underlined numbers in parentheses refer to 
items in the list of references at the end of this paper. 
Work conducted for the Bureau under contract. 



Projects listed in table 1 range from 
minor efforts (such as providing an im- 
proved technology basis for the American 
National Standards Institute and the 
International Standards Organization to 
use in establishing standards) to major 
multiphase projects involving up to 
five supporting industry and Government 
service contracts (such as the fatigue 
study). The research on low-frequency 
vibrations is a TCRC technical assistance 
effort for the Office of Surface Mining 
on blasting vibrations above abandoned 
underground mine workings. 

Blasting Safety 

Research on blasting safety at TCRC is 
concerned with safer blasting practices 
and blast designs primarily for surface 
mines and underground raetal-nonmetal 
mines. The Bureau also conducts explo- 
sives research at the Pittsburgh Research 



Center, which focuses on the safe use and 
evaluation of permissible explosives and 
permissible blasting methods for under- 
ground coal mines, and research on the 
properties and explosives chemistry re- 
lated to safe explosive performance, 
storage, and transportation. 

Table 2 summarizes the major blasting 
projects at TCRC in the Bureau's Health 
and Safety Program. Initially, all work 
was contracted out. More recently, the 
TCRC has conducted in-house research on a 
wide range of significant problems re- 
lated to safer blasting procedures. 
Examples of this in-house research are 
blast designs for safer and more stable 
highwalls and the development of materi- 
als for blasters training. 

Research on blasting safety is guided 
by the analysis of blasting accident 
statistics. This analysis now covers 
8 yr, and has determined the most fre- 
quent causes of blasting accidents and 



TABLE 2. - Research at TCRC since 1980 on blasting safety 



Research area 



Fiscal 
years 



Key researchers 



Significant 
publications 1 



Certification of blasters 

Blasting manual 

Misfires 

Blast area security 

Highwall stability 

Blasters training 

Flyrock control 

Blasting accident 
analysis. 
'Only the senior author 
items in the list of refe 

9 

Work conducted for the 



1978 

1980-81 
1982 

1981-84 

1983-84 

1984-86 

1986 
1981- 
(ongoing) 



E. I. DuPont de Nemours 

& Co. 2 

R. A. Dick , 

L. R. Fletcher , 

Mining & Marketing 

Associates. 2 
V. J. Stachura, 

L. R. Fletcher. 
L. R. Fletcher and 

M. A. Peltier. 

L. R. Fletcher , 

D. V. D'Andrea , 



Coulson and Southall 

(10). 
Dick, IC 8925 (1_1). 
Fletcher (12). 
Bennett (13-14). 

Stachura, RI 8916 

0_5), RI 9008 (16). 
In preparation. 

Fletcher (17). 
D'Andrea (_18). 
Peltier (19). 



is listed here. Underlined numbers 
rences at the end of this paper. 
Bureau under contract. 



in parentheses refer to 



identified where the most hazardous situ- 
ations exist. Through this analysis, re- 
searchers have been able to determine the 
most critical industry needs and address 



these needs through research efforts in 
areas such as blast area security, mis- 
fires, and flyrock control. 



SUMMARY 



The blasting research effort at TCRC is 
concerned with improved productivity, 
blasting vibrations, and blasting safety. 
The following papers summarize research 
projects on surface mine blasting that 
have been carried out and reported on 
since the Bureau last held a Technology 
Transfer seminar on blasting in December 
1980. 

Highlights of the TCRC research include 
evaluations and recommendations for vi- 
bration measurement methods near build- 
ings and a comprehensive study of a test 



house showing that low-level vibrations 
from repeated blasting did not damage 
structures. Blast design studies found 
that the best fragmentation was achieved 
when delays between blastholes were at 
least 1 ms/ft of burden. Highwall sta- 
bility was improved using longer periods 
in the row of blastholes that formed the 
highwall. Accident analysis indicated 
that failure of the blast area security 
system is the major cause of blasting 
accidents. 



REFERENCES 



1. Robertson, D. A., J. A. Gould, 
J. A. Straw, and M. A. Dayton. Survey 
of Blasting Effects on Ground Water Sup- 
plies in Appalachia. Volume I (contract 
J0285029, Philip R. Berger and Associ- 
ates, Inc.). BuMines 0FR8(l)-82, 1980, 
159 pp. 

2. Stachura, V. J. , D. E. Siskind, and 
J. W. Kopp. Airblast and Ground Vibra- 
tion Generation and Propagation From 



Contour Mine Blasting. BuMines RI 8892, 
1984, 31 pp. 

3. Acoustical Society of America (New 
York). Guide to the Evaluation of Human 
Exposure to Vibrations in Buildings. 
American National Standards Institute 
(ANSI) S3. 29. 1983, 9 pp. 

4. Stagg, M. S., D. E. Siskind, M. G. 
Stevens, and C. H. Dowding. Effects of 



Repeated Blasting on a Wood-Frame House. 
BuMines RI 8896, 1984, 82 pp. 

5. Siskind, D. E. , and M. S. Stagg. 
Blast Vibration Damage to Structures. 
Paper in Proceedings 16th Annual Insti- 
tute on Coal Mining Health, Safety, and 
Research. VA Polytech. Inst. , Dep. Min. 
and Miner. Eng. , Blacksburg, VA, 1985, 
pp. 199-213. 

6. Woodward, K. A. , and F. Rankin. 
Behavior of Concrete Block Masonary Walls 
Subjected to Repeated Cyclic Displace- 
ments. NBS (Tech. Rep.) NBSIR 83-2780, 
1983, 178 pp. 

7. Siskind, D. E., and M. S. Stagg. 
Blast Vibration Measurements Near and On 
Structure Foundations. BuMines RI 8969, 
1985, 20 pp. 

8. Kopp, J. W. , and D. E. Siskind. 
Effects of Millisecond-Delay Intervals on 
Vibration and Airblast From Surface Coal 
Mine Blasting. BuMines RI 9026, 1986, 
44 pp. 

9. Nutting, M. J., M. S. Stagg, and 
D. R. Barlet. Effects of Delay Time on 
Fragmentation. Paper in Rock Mechanics: 
Key to Energy Production, (Proc. 27th 
U.S. Symp. on Rock Mechanics, ed. by 
H. C. Hartman Univ. AL, Tuscaloosa, AL, 
June 23-25, 1986). Soc. Min. Eng., 1986, 
pp. 449-454. 

10. Coulson, J. R. , and L. T. South- 
all, II. Considerations for the Certifi- 
cation of Blasters (contract J0285012, 
E.I. DuPont de Nemours & Co., Inc.). Bu- 
Mines OFR 59-81, 1980, 77 pp.; NTIS PB 
81-214116. 

11. Dick, R. A., L. R. Fletcher, and 
D. V. D'Andrea. Explosives and Blasting 
Procedures Manual. BuMines IC 8925, 
1983, 105 pp. 

12. Fletcher, L. R. , and D. V. 
D'Andrea. A Study of Misfires in Min- 
ing. Paper in Proceedings of the 9th 
Conference on Explosives and Blasting 
Technique, ed. by C. J. Konya. (Dallas, 



TX, Jan. 31-Feb. 4, 1983). Soc. Explos. 
Eng., Montville, OH, 1983, pp. 123-132. 

13. Bennett, J. Survey of Safety Pro- 
cedures for Guarding Blast Affected Areas 
(contract J0205019, Mining & Marketing 
Associates, Inc.). BuMines OFR 88-82, 
1981, 74 pp.; NTIS PB 82-246042. 

14. D'Andrea, D. V., and J. Bennett. 
Safeguarding of Blast-Affected Areas. 
Paper in Proceedings of the 10th Confer- 
ence on Explosives and Blasting Tech- 
nique, ed. by C. J. Konya (Lake Buena 
Vista, FL, Jan. 29-Feb. 2, 1984). Soc. 
Explos. Eng., Montville, OH, 1984, 
pp. 110-119. 

15. Stachura, V. J., and L. R. Flet- 
cher. Delayed Blasting Tests To Improve 
Highwall Stability — A Progress Report. 
BuMines RI 8916, 1984, 24 pp. 

16. Stachura, V. J. , L. R. Fletcher, 
and M. A. Peltier. Delayed Blasting 
Tests To Improve Highwall Stability — A 
Final Report. BuMines RI 9008, 1986, 
12 pp. 

17. Fletcher, L. R. , and D. V. 
D'Andrea. Control of Flyrock in Blast- 
ing. Paper in Proceedings of the 12th 
Conference on Explosives and Blasting 
Technique, ed. by C. J. Konya (Atlanta, 
GA, Feb. 9-14, 1986). Soc. Explos. Eng., 
Montville, OH, 1986, pp. 167-177. 

18. D'Andrea, D. V., and L. R. Flet- 
cher. Analysis of Recent Mine Blasting 
Accidents. Paper in Proceedings of the 
9th Conference on Explosives and Blasting 
Technique, ed. by C. J. Konya (Dallas, 
TX, Jan. 31-Feb. 4, 1983). Soc. Expl. 
Eng., Montville, OH, 1983, pp. 105-122. 

19. Peltier, M. A., L. R. Fletcher, 
and D. V. D'Andrea. Coal Mine Blasting 
Accidents. Paper in Symposium on Engi- 
neering Health and Safety in Coal Mining, 
ed. by A. Wahab Khair (New Orleans, LA, 
Mar. 2, 1986). Soc. Min. Eng. AIME, 
1986, pp. 174-181. 



REDUCING ACCIDENTS THROUGH IMPROVED BLASTING SAFETY 
By Larry R. Fletcher 1 and Dennis V. D'Andrea 2 



ABSTRACT 



The Bureau of Mines investigated three 
of the major causes of mine blasting ac- 
cidents: inadequate blast area security, 
excessive flyrock, and misfires. 

Accidents resulting from inadequate 
blast area security occur during sched- 
uled blasting because of failure to clear 
the blast zone, inadequate guarding, and 
failure of personnel to follow instruc- 
tions, retreat to a safe location, and/or 
take adequate cover. 

Excessive flyrock is produced when 
there is too much explosive energy for 
the amount of burden, stemming is inade- 
quate, or the explosive energy is too 
rapidly vented through a zone of weakness 
in the rock. Geology, improper blast de- 
sign, or carelessness can cause unwanted 



flyrock. The operator must change blast- 
ing methods when shooting in geology that 
favors the production of flyrock. 

Most misfire accidents are caused by 
drilling into bootlegs in underground 
metal and nonmetal mines. Improper dis- 
posing of misfires is the second most 
frequent cause of misfire accidents, and 
some accidents are due to impact initia- 
tion of explosives in the muckpile. Mis- 
fires are usually caused by misunder- 
standing, improper use, or some failure 
of the initiation system. Other causes 
are cutoffs, insufficient firing current, 
inadequate priming, improper explosive 
storage, and damage to the initiation 
system. 



INTRODUCTION 



Analysis of mine blasting accidents 
shows that the five most frequent causes 
of accidents are premature blasts, in- 
adequate blast area security, excessive 
Elyrock, misfires, and fumes. Over 70 
pet of all blasting accident injuries 
from 1978 through 1985 have been caused 
by three of these: inadequate blast area 
security, flyrock, and misfires, all 
of which are discussed in this paper, 
based on earlier Bureau of Mines research 

a-2). 3 

Each mining operation has a normal fly- 
rock range, the distance from the blast 
at which flyrock can be expected, based 
on blasting experience at that operation. 
The distance of the normal flyrock range 
will vary from a few feet in an area 
strip coal mine blast to more than a mile 

^Mining engineering technician. 

^Research supervisor. 
Twin C ities Research C enter, Bureau of 
Mines, Minneapolis, MN. 

■^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



with poorly controlled shots. A safety 
factor is added to the normal flyrock 
range to determine the blast area to be 
cleared and secured before the blast is 
detonated. Rock that travels beyond the 
secured blast area is excessive flyrock 
(fig. 1). A distinction is made between 
injuries that occur within the estab- 
lished blast area and injuries that are 
the result of excessive flyrock projected 
beyond the blast area. 

Anyone who remains in the blast area, 
such as the shot firer, must have ade- 
quate protection from flyrock such as 
that provided inside a blast shelter. 
The shot firer is frequently guilty of 



Excessive flyrock 
< 



Excessive flyrock 
> 



Secured blast area 



, Normal flyrock range 
< '- ^— > 



Blast site 
< > 



y>/yvv vy y'vs'vy v 

A, <. "-' ' ' / > v 



-L -t-' 



< Vs' " 



FIGURE 1.— Blast site, normal flyrock range, secured blast 
area, and excessive flyrock region for surface mine blast. 



shooting and observing blasts from within 
the blast area without adequate protec- 
tion. When a shot firer pays inadequate 
attention to the hazards of flyrock, a 
permissive attitude toward guarding and 
protecting the blast area is created. 

A misfire results when explosives fail 
to detonate as planned during a mine 
blast. It is difficult, if not impossi- 
ble, to determine how frequently misfires 
occur. When operators are asked, they 
usually reply, "Rarely, if ever, do we 
have a misfire." The truth is, however, 
that misfires are fairly common, but 



because many people feel that misfires 
reflect the quality of their work, they 
are reluctant to report them. In addi- 
tion, the increased use of non-cap-sen- 
sitive blasting agents with lower shock 
sensitivity has generated an attitude of 
indifference about misfires at some oper- 
ations. Because of these two factors, it 
is rare that a misfire that does not in- 
volve an injury is ever reported. 

This paper discusses the elements of 
effective blast guarding, the causes and 
control of flyrock, and the causes, de- 
tection, and disposal of misfires. 



BLAST AREA SECURITY 



The blast area security system is the 
means by which a mine operator prevents 
injury to people or damage to equipment 
when a scheduled blast is detonated. 
Most blasting accidents in surface and 
underground mines occur during scheduled 
blasting and are due to inadequate blast- 
site security. The result is often un- 
necessary injury and/or death, caused, in 
most cases, by flying rock. Mine person- 
nel, visitors, and even trespassers can 
inadvertently wander into a blasting zone 
unless proper procedures exist to ensure 
that all personnel are cleared from the 
blast zone and kept safely away until af- 
ter blasting is completed. 

GENERAL REQUIREMENTS FOR BLAST GUARDING 

The basic requirements of a blast-site 
security system at either a surface or 
underground mine are (1) to move person- 
nel and equipment out of range of flyrock 
from the blast and (2) to prevent move- 
ment of personnel back into the blast 
zone. This includes visitors and tres- 
passers, along with mine personnel. 

The most effective procedures for ac- 
complishing blast-site security objec- 
tives are to — 

Have blasting personnel physically 
clear the blast-affected zone; 

Account for personnel, to establish 
that no one is present in the blast zone 
at blast time; and 

Place guards beyond the blast zone, at 
entries leading into the blast site, to 



keep personnel from moving into the blast 
zone. 

Elements contributing to the accom- 
plishment of the blast-site security sys- 
tem requirements are as follows: 

Management commitment to safety. 
Training. 

Definition of blasting authority. 
Planning. 

Blast zone boundaries. 
Selection and placement of guards. 
Clearing procedure. 
Location of blast initiation sice. 
Blasting time. 
Blast signals. 
Personnel accounting. 

Communication of blast location and 
time. 

Blasting crew communications. 

Each of these elements is discussed 
below. 

MANAGEMENT COMMITMENT TO SAFETY 

The mines that have developed the best 
blast-guarding systems appear to be those 
in which management has made a definite 
commitment to safety. It has been recog- 
nized at those mines that it is manage- 
ment's responsibility to provide raining 
personnel with a systematic, safe blast- 
guarding method and to provide sufficient 
training to allow mining personnel to 
operate intelligently within the estab- 
lished system. 



Blast-site security objectives are gen- 
erally accomplished through the develop- 
ment of a blast-guarding standard operat- 
ing procedure specifically designed to 
fit the requirements of a given mine. 

TRAINING 

Having developed a standard blast- 
guarding procedure, mine management is 
obligated to inform its employees of 
their responsibilities. Operations with 
good blasting security systems utilize 
information regarding proper blast guard- 
ing as training material for new hires 
and for retraining experienced miners. 
It is important that blast guarding be 
treated specifically and in detail as a 
real safety problem. 

DEFINITION OF BLASTING AUTHORITY 

One experienced individual must be re- 
sponsible for the blast. Whether this 
person is a blasting superintendent, 
foreman, or crew leader, it is extremely 
important that one key individual direct 
the blast-guarding process for any given 
blast. 

PLANNING 

Preblast planning is an important as- 
pect of blast guarding. Mines that have 
standardized their blast-guarding methods 
reduce the requirement for daily planning 
by making certain blast-associated deci- 
sions routine. This is particularly true 
at mines that have adopted central blast- 
ing systems and consistent methods for 
identifying crew member locations before 
a blast. 

BLAST ZONE BOUNDARIES 

The boundary around the blast beyond 
which personnel and equipment will be 
safe from flyrock must be determined. 
Guard posts must be located outside that 
boundary at all entries into the blast 
zone. Flyrock range, of course, is de- 
pendent on a number of factors including 
blast configuration, rock hardness, stem- 
ming, geology, etc., and is determined 
for each particular location and blast. 



SELECTION AND PLACEMENT OF GUARDS 

It is the general consensus throughout 
the industry that human guards should be 
used to guard blast-site entries whenever 
possible. Entrances to blast areas are 
sometimes left unguarded because it is 
either impractical or unsafe to place a 
person on post. When an entry is un- 
guarded, a barricade or sign is used to 
warn people, but these methods are gener- 
ally considered inadequate. 

Persons selected as guards should be 
sufficiently trained to understand the 
serious responsibility they have to pre- 
vent any movement of personnel back into 
the blast zone. The absolute authority 
of the guard to prevent movement beyond 
the post must be established. The guards 
themselves must be placed in a location 
safe from flyrock and fumes. 

CLEARING PROCEDURE 

Mine management should establish a def- 
inite clearing procedure as part of the 
blast-guarding standard operating proce- 
dure. The clearing path will depend on 
mine layout, and considerable variance 
can exist within a given mine. Clearing 
personnel in surface mines are assigned 
to drive or walk through the blast zone 
prior to the blast. In mines where low- 
mobility equipment such as drills and 
shovels are left in the blast zone, this 
equipment should be inspected to be sure 
all personnel have left. 

Clearing procedures in underground 
mines vary considerably. Since blasters 
frequently finish loading underground 
rounds well before the scheduled shift- 
change blast time, last-minute clearing 
must be done. If effective clearing pri- 
or to blasting is impractical, it is es- 
sential to use personnel accounting and 
keep miners informed of the time and lo- 
cation of the blast. Blast-guarding sys- 
tems generally benefit from the inherent 
redundancies associated with requiring 
both clearing and personnel accounting in 
the blast zone. 



LOCATION OF BLAST INITIATION SITE 

A definite initiation site (blasting 
location) should be designated for all 
blasting areas in the mine or pit. Loca- 
tion of the blast initiation site should 
be based on a conscious management deci- 
sion as part of the preblast planning 
process. 

The purpose of designating location is 
twofold. First, it provides management 
with a means of fulfilling its respon- 
sibility to ensure that the blaster re- 
treats to a safe location before blast- 
ing. The recent Bureau analysis of 
blasting accidents revealed that many 
surface and underground accidents were 
caused by unsafe location of the shot 
firer. Second, the location of a desig- 
nated blast initiation station can easily 
be communicated to all mine personnel as 
the required checkpoint for personnel ac- 
counting when persons are entering or 
leaving that area of the mine or pit. 

Minimum recommended requirements for a 
blasting site are as follows: 

The site should be located outside the 
blast zone, protected from flyrock and 
upwind of blast fumes. Blast initiation 
from inside the blast zone should be per- 
mitted only if a secure blast shelter is 
used. 

The site should be equipped with a 
highly visible sign giving the blast 
time. The sign should also notify visi- 
tors that they are required to check in 
and out at that station. 

BLASTING TIME 



done concurrently 
blast. 



with the production 



Predictability is an 
in a process with the 
tial of blasting. The 
ommended that efforts 
lish consistent blast 
blast is initiated at 
end of a shift, blast 
be strictly controlle 
Timing for secondary 
cording to the needs 
mine. Many mines, bo 
derground, have deve 
which secondary bias 



essential element 
destructive poten- 

refore, it is rec- 
be made to estab- 

times. Whether the 
lunch time or the 
scheduling should 

d where possible. 

blasting varies ac- 
of the particular 

th surface and un- 

loped systems in 

ting is regularly 



BLASTING SIGNALS 

Audible signs are used at most surface 
mines to warn of the impending blast. In 
some cases, an air horn or electric siren 
is mounted at a fixed power source. In 
others, particularly when the pit is 
large, sirens or horns are mounted on 
vehicles used for clearing the blast 
zone. In either case, the devices used 
should be loud enough to be heard by per- 
sonnel approaching the blast zone from 
any direction. In addition, considera- 
tion should be given to prevailing wind 
directions and velocities. Many surface 
mines use flashing lights on blast-asso- 
ciated vehicles as a supplement to sirens 
and horns. 

The duration and pattern of audible 
signals used at the mines vary consid- 
erably. Some start 10 min prior to the 
blast, while others start 2 min before 
the blast. Some signals continue 
throughout the blast, while others are 
interrupted during the blast. A second 
signal after the blast signifies that all 
is clear. 

The signal must fulfill three require- 
ments, as follows: 

It must be loud enough to warn person- 
nel inside and near the zone of the im- 
pending blast. 

In the event that the clearing process 
has failed and persons remain exposed to 
the impending blast, the signal should 
allow sufficient time for them to take 
cover or leave the area. 

It must clearly indicate when it is 
safe to reenter the blast zone. 

The air-powered whistle used at the 
Union Carbide Creek Pine Creek Mine (fig. 
2) is a simple and effective means of 
warning that a blast is about to be ini- 
tiated, particularly in areas where en- 
tries are left unguarded. A whistle im- 
proves blasting-site security when used 
as a backup to human guards and is a very 
cost-effective device. Easy to use, the 
whistle is simply screwed onto a drop on 
the air line and turned on as the blaster 



10 




°/ 8 in NPT \ ^ 46" in schedule 40 pipe \ 

Notch and weld Weld plug _\ 

reed in place in end 

FIGURE 2.— Air whistle used at Pine Creek Mine (NPT— National 
Pipe Taper). 



retreats from the face to begin the 
clearing process. It remains on through- 
out the blasting procedure and is turned 
off by the first person entering the area 
after the blast. 

PERSONNEL ACCOUNTING 

Personnel accounting refers to deter- 
mination of the locations of personnel 
throughout the mine site. In essence, it 
is a tracking system and should be de- 
signed, at the very least, to account for 
the presence of personnel within the zone 
of a blast. Whenever reasonably possi- 
ble, accounting stations should be estab- 
lised near all blast zones, preferably at 
the locations from which the blasts are 
regularly initiated. At a minimum, an 
effective personnel accounting system 
must establish that no personnel are oc- 
cupying the blast zone at blast time. 
The personnel accounting system is an im- 
portant backup to the clearing process. 
However, some large open pit operations 
that blast several times each day rely on 
clearing as the only practical method of 
ensuring that there are no people in the 
blast area. When clearing is the only 
method used, the clearing personnel must 
conscientiously cover the entire blast- 
affected area prior to every blast. 

Many mines account for personnel by 
means of a head count at some safe loca- 
tion prior to the blast. While a head 
count is better than not accounting for 
personnel at all, it allows considerable 
potential for error. 

The system of using a blasting board is 
applicable at many underground and sur- 
face mines. Blasting boards are placed 
at safe locations where miners congregate 



during the blast. The board contains a 
tag for every person assigned to a par- 
ticular station plus tags for visitors. 
When blast time arrives, supervisors at 
each safe location merely need to look at 
the board to ensure that all personnel 
have returned from the immediate blast 
zone. 

Surface mines seem to place less empha- 
sis on personnel accounting than do the 
underground mines. This is somewhat jus- 
tified by the fact that clearing proce- 
dures are more effectively supervised and 
performed within surface blast areas be- 
cause of better visibility. It is also 
impractical for some large surface mines 
that blast several times each day to ac- 
count for all mine personnel before each 
blast. Fortunately, most of these mines 
have a limited flyrock range and can re- 
liably use the clearing process. How- 
ever, where practical, personnel account- 
ing should act as a backup system to the 
clearing process. Blasting safety at 
many surface mines would benefit from an 
improvement in personnel accounting. 

COMMUNICATION OF BLAST 
LOCATION AND TIME 

It is important that all personnel en- 
tering a mine be aware of the exact loca- 
tion of blasting sites and the schedule 
for blast initiation. Surface and under- 
ground mines that blast on a daily basis 
can list the blasting times on signs at 
appropriate entrances into the mining 
areas. Since blast location changes fre- 
quently at most mines, mine personnel 
should be informed on a daily basis of 
the exact locations of all scheduled 
blasts. 

Notices regarding blast-site locations 
in underground mines should be placed at 
the personnel accounting station for the 
blasting area. As personnel check in at 
the blasting station, they are informed 
of the blast time and location. This 
particularly benefits personnel who are 
not regularly assigned to a section for 
which blasting is scheduled but who are 
required to move throughout the mine. 



11 



BLASTING CREW COMMUNICATIONS 

Some surface mines utilize hand signals 
to communicate between the shot firer and 
blast guards. However, hand signals do 
not provide the positive, instantaneous 
communication so necessary to blasting- 
site security. A hand signal will not 
always immediately attract the blaster's 
attention. Poor visibility, a blaster's 
not wearing corrective eyeglasses, or sun 
shining in the blaster's eyes are all 
cases in which a hand signal might not be 
easily seen. 

Guards should use warning devices 
capable of quickly and positively gaining 
the blaster's attention. The two-way 
radio does this best. Safe separation 
distances between transmitters and elec- 
tric blasting circuits can be determined 



by reference to the Institute of Makers 
of Explosive (IME) Publication 20, "Safe- 
ty Guide for the Prevention of Radio Fre- 
quency Radiation Hazards in the Use of 
Electric Blasting Caps" (4_) . Another 
source is the Du Pont "Blaster's Hand- 
book" (_5 ) . Radio communication for blast 
guarding is safe if the IME guidelines 
are followed. Transmitters should be 
limited to 5 W or less, and a boundary 
should be defined around the blast pat- 
tern within which the use of radios is 
forbidden. 

An alternative is the use of nonelec- 
tric blast initiation systems, which 
are widely used in the mining industry. 
These systems cannot be initiated by 
radio frequency energy and are com- 
pletely compatible with close radio 
communication. 



FLYROCK 



Excessive flyrock is rock that is pro- 
jected beyond the normal blast-affected 
area. It is generated when there is too 
much explosive energy for the amount of 
burden, when stemming in insufficient, 
or when the explosive energy is rapidly 
vented through a plane of weakness. Ex- 
cessive flyrock is responsible for 24 pet 
of the blasting accident injuries that 
occur in surface mining (1978-85). Ex- 
cessive flyrock can be the result of 
blast site geology and/or rock condi- 
tions, improper blast design, or care- 
lessness. Flyrock control is achieved by 
careful attention to blast design, blast- 
site inspection, blasthole layout, blast- 
hole drilling, and blasthole loading 
practices. 

FLYROCK CAUSES 

Geology and rock conditions can cause 
the generation of flyrock. Geologic fea- 
tures such as mud seams, natural joint or 
bedding planes, fractures, or cavities in 
the rock can result in excessive flyrock. 
Mud seams and fractures are planes of 
weakness through which explosive gases 
can rapidly vent and accelerate rock 
fragments. Cavities can accidentally 



become filled with too much explosive for 
the amount of rock burden, resulting in 
large flyrock distances. Blastholes can 
penetrate openings from abandoned under- 
ground mines to create a dangerous con- 
dition similar to that resulting from 
natural cavities. Fracturing due to 
backbreak or overbreak from previous 
blasting can also cause dangerous planes 
of weakness. A ragged highwall face or 
an overhang can result in diminished bur- 
den along the front row of holes. 

Excessive flyrock can be generated if 
blasts are not properly designed. Any 
blast design feature that results in in- 
sufficient explosive confinement or the 
rapid venting of the explosive gases can 
create a problem. Blast design errors 
such as too high a powder factor, an in- 
adequate burden, too short a stemming re- 
gion, failure to use stemming, improper 
delays between rows, or the wrong blast- 
hole delay sequence can result in un- 
wanted flyrock. The wrong delay sequence 
can cause diminished burden if the delay 
is too long. Cratering and blowouts can 
occur when back holes fire before front 
holes. A very short delay can result in 
too much confinement and, again, crater- 
ing and blowouts. 



12 



Unfortunately, carelessness is a lead- 
ing cause of excessive flyrock. Care- 
lessness during any part of the blast de- 
sign, the blasthole pattern layout, the 
drilling of the blastholes, the loading 
of the blastholes, or the hookup of the 
initiation system can create a dangerous 
situation. The loading of explosives too 
near the collar of the blasthole is a 
common cause of flyrock. 

CONTROL OF FLYROCK 

The control of flyrock starts with 
proper blast design. The correct burden 
must be used. A small burden will not 
contain the explosive energy, while using 
a large burden may result in cratering 
and/or blowouts. The bench height, bur- 
den, and stemming region must be such 
that the blasted rock movement is primar- 
ily horizontal and outward, and not up- 
ward. In multiple-row shots, the delay 
between rows must be long enough to allow 
rock from an earlier row to move out so 
that the next row will have adequate re- 
lief. Insufficient relief can cause fly- 
rock. However, the delay must not be so 
long that cutoffs occur and cause mis- 
fires that increase the burden on later 
firing holes, again resulting in blowouts 
and flyrock. 

In designing a blast, relationships be- 
tween charge diameter, burden, spacing, 
subdrilling, stemming region, and bench 
height are available for initial esti- 
mates (6). The type of explosive, the 
priming, and the initiation system must 
be selected. Toe priming reduces fly- 
rock, compared with collar priming. 
Decisions must be made on the type of 
blasthole pattern, square or staggered, 
and on the delay sequence. The powder 
factor is calculated to ensure that the 
quantity of explosive being used is with- 
in the range of that normally used in 
surface mine blasting. These initial ap- 
proximations must be modified for the 
particular blasting situation and may be 
further modified after experience with a 
number of blasts. 

Before the blasthole pattern is laid 
out in preparation for drilling, a care- 
ful inspection of the blast site should 
be made. The face should be examined for 



ruggedness, overhangs, fractures, zones 
of varying competence, and amount of toe 
burden. The blast site should also be 
inspected for backbreak, jointing, mud 
seams, voids, and other zones of weak- 
ness. Any of these blast-site features 
could cause excessive flyrock. 

The layout of the blasthole pattern 
starts with the front row. If the ver- 
tical face has overhangs, is concave, or 
is irregular, the burden at some point 
may be reduced and violent cratering 
could occur. Faces with backbreak, open 
joints, weak zones, and mud seams will 
allow rapid venting of explosion gases 
with flyrock. In addition, the burden 
will not pull (be removed) as planned, 
causing an increase in the burden for 
later holes, which results in cratering 
at the top of the bench. If the face is 
sloped, the toe burden will be larger 
than the crest burden unless angled holes 
are used. If the normal column load is 
used when there is a sloped face, there 
will be flyrock because of the short 
crest burden. Also, the toe may not 
pull, producing a buildup in front of 
later holes, again resulting in flyrock. 
Adjustments in hole locations and powder 
columns in the front row should be made 
when conditions exist near the face that 
favor the generation of flyrock. Once 
the front row is established, the bal- 
ance of the shot can be laid out. A tape 
should be used to ensure accurate spacing 
and burden distances. 

Locating a blasthole close to an open 
fracture will provide a weak zone. The 
shot will break into the fracture, vent 
with flyrock, and produce poor fragmenta- 
tion. The same kind of venting can occur 
when a hole is abandoned and a second 
hole is drilled a few feet away. To pre- 
vent this, the first hole should be back- 
filled. Where open fractures are pres- 
ent, they can be backfilled, but this is 
difficult and time consuming. The best 
way to handle fractures from previous 
blasting is to eliminate the cause of 
overbreak. 

Accurate drilling is essential. The 
holes must be located in accordance with 
the blast design and drilled at the cor- 
rect angle and to the proper depth. With 
a high face and smaller diameter holes, 



13 



extra care should be taken to ensure that 
holes are drilled at the proper angle. 
Deviated blastholes can result in burdens 
much smaller or larger than planned. The 
driller should provide a log of each hole 
that includes depth drilled, problem 
drilling zones, and any changes in pene- 
tration rate. This information could in- 
dicate voids or zones of weakness. 

All holes must be checked before explo- 
sives are loaded, to ensure that their 
location and depth are in accordance with 
the blast design and the driller's log. 
If blocked holes are undetected, the ex- 
plosive column can be loaded too near the 
collar, leaving an insufficient stemming 
region. Partially blocked holes can also 
be a problem and may not be detected with 
a weighted tape. On sunny days, a mirror 
may be used to check for obstructions. 
Before an attempt is made to load a short 
hole or a partially blocked hole, correc- 
tive action such as redrilling, changing 
the planned explosive load, or abandoning 
and backfilling of the hole should be 
considered. 

Open joints and cracks can extend into 
a blasthole. The presence of cracks can 
often be detected by a lack of drill cut- 
tings at the top of the hole. However, 
unless corrective action is taken during 
the loading of bulk products, these open 
fractures can be loaded with explosives, 
resulting in excessive energy for the 
hole. There will be less confinement of 
the explosive charge. This may affect 
the performance of some explosive prod- 
ucts as well as provide a weak zone that 
may blow out. 

Frequent checks of the powder-column 
rise during loading are important. A 
slower than normal rise may indicate a 
void, while a sudden rise indicates a 
blockage. When a blockage occurs during 
explosive loading, care must be taken to 
ensure that the explosive load above the 
blockage will detonate. This will re- 
quire the placing of another primer and 
detonator in the column above the block- 
age. When a void is indicated, loading 
must be stopped. If the void is not un- 
duly large, backfilling with stemming 
material will correct the problem. 
Another primer must be added and loading 



continued. If the void is too large for 
this to be practical, the hole may have 
to be abandoned and another hole drilled 
nearby. Redrilling should not be done if 
there is a possibility of drilling into 
explosives. In some cases, the hole can 
be plugged just above the void. Once a 
plug is formed, the explosive loading 
can resume. In operations where voids 
are common, a special system for borehole 
plugging should be developed. 

Checking the column rise will prevent 
accidental overloading of the blasthole. 
Maintaining sufficient stemming is an 
important factor in flyrock control. 
Stemming lengths of 0.7 to 1.0 times the 
burden are commonly used. When collar 
priming is used, the stemming length may 
need to be increased because of the 
greater potential for violence with top 
priming. Crushed and sized rock is the 
best material to use for stemming, but 
drill cuttings are commonly used because 
of availability and economy. Large 
pieces of rock or other material should 
never be mixed with the stemming as they 
can become missiles if there is a blow- 
out. Large rocks can also cut off or 
damage the initiation system and cause a 
misfire. 

Care must be taken to ensure that the 
initiation system is properly hooked up 
and that the delays are correct. A final 
check of the hookup is imperative. 

The secondary blasting of boulders too 
large for the loading equipment or crush- 
er is required at some operations. Sec- 
ondary blasting can produce dangerous 
flyrock even though the charge? are 
small. Determining the blast area for 
this kind of shooting is very difficult, 
and careful clearing and guarding are re- 
quired. Although secondary blasting is 
frequently done on-shift and as needed, 
it is best to shoot at a standard time 
such as during shift change. Second 
blasting can be done at the same time as 
primary blasting if the same immediate 
area is involved. However, if the blasts 
are widely separated, there will be two 
sources of flyrock to guard against. 

The reshooting of misfired blastholes 
can generate dangerous flyrock when there 
is reduced burden and reduced confinement 



14 



of the explosive charge. When reduced 
burden, distance can exceed the normal 
flyrock range, from production blasts, 
and a larger-than-normal blast area is 



required, which must be cleared and 
guarded. It is best to shoot misfired 
blastholes during shift change or at the 
same time as primary blasts. 



MISFIRES 



A misfire results when explosives fail 
to detonate as planned during a blast. 
A misfire has two basic effects on an 
operation: the safety hazard it presents 
and the increase in mining costs. 

EFFECTS OF MISFIRES 

To most people, a misfire primarily 
represents a safety hazard. With the in- 
creasing use of non-cap-sensitive blast- 
ing agents, the possibility of accidental 
initiation is reduced. However, based on 
U.S. Mine Safety and Health Administra- 
tion (MSHA) data, there are still many 
injuries sustained because of misfire ac- 
cidents. In an 8-yr period (1978-85), 56 
misfire accidents resulted in 63 injuries 
and 6 fatalities. The majority of mis- 
fire accidents (75 pet) occur in under- 
ground mines, with 54 pet in underground 
metal-nonmetal mines. These numbers are 
not surprising because it is not the 
total amount of explosives used, but the 
number of shots or holes fired that 
provides the opportunity for misfires. 
Also, underground mines use smaller diam- 
eter charges, fired on smaller spacings, 
which are prone to misfire. In addition, 
visibility in underground mines is gener- 
ally poor, so detection of misfires is 
hampered. 

The remaining 25 pet of misfire acci- 
dents occurred in surface mines, with 11 
pet in surface coal mines and 14 pet in 
surface metal-nonmetal mines. 

Misfires that result in an accident 
with injuries, fatalities, and/or equip- 
ment damage involve obvious costs; how- 
ever, there are other costs due to mis- 
fires that are not obvious. Because the 
effects of misfires will vary greatly, 
operators should conduct probable cost 
analyses of misfires at their mines. 
Those who do will place greater emphasis 
on avoiding misfires. 

In addition to the cost of disposal of 
the misfire, there are the direct costs 



in additional drilling, explosives, prim- 
ers, detonators, and labor. In some un- 
derground mines, the failure of a "cut" 
hole could result in the loss of the en- 
tire round. 

The handling of boulders that require 
secondary breakage is another cost fac- 
tor. Increased digging time and greater 
wear and damage to equipment (especially 
bucket teeth) result in lower productiv- 
ity and higher maintenance costs. Mis- 
fires frequently are the cause of high 
bottom, which results in reduced produc- 
tion and higher maintenance costs on 
mucking equipment. In addition, haulage 
vehicles traveling over rough terrain 
will increase the haulage cost and vehi- 
cle maintenance. Often, these humps must 
be drilled and blasted, which constitutes 
another cost. 

The misfiring of one hole will increase 
the burden on a later hole, causing cra- 
tering with excessive flyrock and over- 
break. Flyrock is a leading cause of 
personal injury and equipment damage. 
Overbreak may extend beyond the burden 
for the first row of holes for the next 
shot, which can cause problems in drill 
setup for the next shot. Overbreak at 
the final pit wall could produce ground 
control problems with very high cost and 
even loss of ore. 

CAUSES AND AVOIDANCE OF MISFIRES 

During the preparation and initiation 
of blasts, there are many aspects that 
may result in misfires. The most fre- 
quently stated cause of misfires is the 
incorrect use of the initiation system, a 
problem common to all initiation systems. 
A major contributing factor is the lack 
of understanding by blasting personnel of 
how the system works. Unless the blast- 
ing personnel have a full understanding 
of the initiation system, even minimal 
changes in a shot can result in poor 
blast performance and misfires. 



15 



Damage to the initiation system or ex- 
plosives column is another common source 
of misfires. Causes include poor work 
practices on the blast site and rock 
movement that produces cutoffs. 

Damage can often occur while stemming 
is shoveled in and when wires, cords, or 
tubes are stepped on. Even driving over 
the initiation systems is not uncommon at 
some mines. The wiring-in or hookup of a 
shot, regardless of the initiation system 
used, is a very important part of blast- 
ing. It is important that each initia- 
tion system be checked after the hookup. 
The method of system checkout depends on 
the system used. All electrical hookups 
should be checked with a blaster' s meter. 
All shots should be checked visually. 
Good housekeeping and neat and consistent 
hookup practices are helpful in accom- 
plishing the system checkout. 

Rock movement may cut off the explo- 
sives columns and result in misfired ex- 
plosives. Uplift, as well as horizontal 
displacement, is a factor in cutoffs. An 
area surface coal mine may have 100 or 
more holes drilled and loaded, which are 
then divided into 4 or 5 shots. Where 
rock movement is generally up, with lit- 
tle or no horizontal movement, the uplift 
can damage loaded holes waiting to be 
wired in for the next shot. 

The proper use of delays is very impor- 
tant in preventing cutoffs. The longer 
the delay between holes, the greater the 
probability of cutoffs. Shorter delay 
times are needed when surface delays are 
used. If longer delay times are required 
for fragmentation and rock displacement, 
in-the-hole delays should be consid- 
ered. Many systems or combinations of 
systems are available to meet delaying 
requirements. 

Geology is another factor that plays 
a major role in causing cutoffs. Frac- 
tures, faults, joints, and bedding planes 
are all zones of weakness that may cause 
burden movement to occur in much shorter 
times than considered normal. The use 
of decking and multiple primers is some- 
times advantageous to avoid cutoffs. 
Rock falling from the walls of the bore- 
holes can cause bridging and explosive 
column separation. 



Hole blockage can also be the result of 
careless work habits, such as knocking 
material into the hole. Rapidly loaded 
cartridged products will often bridge, 
particularly in holes that are partially 
filled with water. Careless work prac- 
tices are often due to the rush placed on 
the blasting crew to get the shot off by 
a given time. 

Poor priming practices frequently cause 
detonation failure. Each product has 
minimum priming requirements. Even when 
proper priming is used, if the primer 
sinks into the mud at the bottom of the 
hole or water enters a hole loaded with 
non-water-resistant products, a misfire 
can occur. Many operators of surface 
mines using larger diameter blastholes 
place the primer at the floor level and 
not at the bottom of the hole in the sub- 
drilled region. In the event of a mis- 
fire, the more sensitive primer and deto- 
nator are more easily retrieved. 

Storage is also an important factor in 
avoiding misfires, since improper storage 
can alter the performance of many prod- 
ucts. The sensitivity of some products 
is dramatically reduced by low-temper- 
ature storage, which can result in mal- 
functioning. Malfunctioning of explo- 
sives products is usually due to improper 
storage rather than quality control prob- 
lems in manufacturing, which are rare. 
Hydrostatic pressure as well as compres- 
sion from the firing of adjacent holes 
has caused some products to lose sensi- 
tivity and misfire. If misfires are to 
be avoided, the blaster must have a com- 
plete understanding of the products and 
conditions under which they can be used. 
This information is available from the 
supplier. 

DETECTION OF MISFIRES 

Sometimes misfires are obvious, some- 
times they are not. Each shot must be 
checked for misfires before mucking is 
begun. When the explosive is lying on 
the muck, detection of a misfire is no 
problem. However, when the explosive is 
buried in broken rock, visual detection 
is unreliable. 



16 



There are a number of clues that may 
indicate a misfire. Most operations have 
standardized blasting practices that have 
fairly uniform results from each shot. A 
change from the norm could indicate mis- 
fires. In surface mines where the shot 
can be observed from a safe location, 
watching and listening to the shot is 
worthwhile. A change in the sound — 
louder or quieter — may indicate misfires. 
Ejection of stemming, cratering, and fly- 
rock may result from too much burden, due 
to misfiring of earlier holes. 

In surface and underground operations, 
the muckpile profile can reveal areas of 
possible misfires. Muck lying mostly to 
one side of the shot, less displacement 
than expected, abnormal backbreak, and 
humps and valleys in the muckpile can all 
be due to misfires. Change in fragmenta- 
tion is a very good indicator of possible 
misfires. Boulders across the top of 
the muckpile are easy to see and could 
indicate misfires. Operators of loading 
equipment should be aware that boulders 
uncovered in lower sections of the pile 
also may be the result of misfires. This 
is common where multiple decks are used. 

Multiple priming can minimize misfires, 
although this method is sometimes not 
used because of additional costs. Con- 
sidering the cost of misfires, perhaps 
multiple priming should be used more fre- 
quently. A double-trunkline or loop sys- 
tem must be used with detonating cord 
systems. Even with the two paths of det- 
onation, all of the cord should be con- 
sumed in the blast. Finding detonating 
cord in the muck is a strong indication 
of a misfire. 

In underground mines, detection is ham- 
pered by poor lighting. It is difficult 
for the loader operator to spot explo- 
sives in the muck. Checks for explosives 
should be made before and during the 
mucking operation. Every bootleg must be 
examined carefully for misfires. A mis- 
fire may have occurred even though there 
is no cap legwire or tubing protruding 
from the hole. Lifter holes are of par- 
ticular concern because they slope down- 
ward and are often water filled, and it 
is easy for loose material to fall into 
them. These conditions promote misfires. 
In addition, lifter holes are the most 



difficult to relocate and check for mis- 
fires. Because of potential misfires, 
blastholes must never be collared in 
bootlegs. 

DISPOSAL OF MISFIRES 

There are two basic methods used to 
dispose of undetonated explosives: to 
recover and destroy the explosives or to 
detonate the misfire in place. Any ex- 
plosive product removed from a misfire is 
considered damaged and must be destroyed 
in a safe manner. The manufacturer is 
the best source of information on de- 
stroying an explosive product. Many 
water-based products do not burn readily 
and are difficult to destroy. The most 
common method is in-place detonation. 
This is good practice in underground 
mines where flyrock is less of a problem. 
However, in many operations, the disposal 
of misfires is the only blasting that is 
done on-shift, and this creates blast 
area security problems. 

When the original initiation system is 
still intact, it can be used to refire 
the charge, except with cap and fuse 
blasting. When a misfire results with 
cap and fuse blasting, the blaster 
should never relight the fuse because it 
may have been shortened, causing unex- 
pected premature initiation, or it may be 
damaged and could produce a hangfire. 
With cap and fuse blasting, repriming is 
essential. 

Misfires are occasionally refired in 
surface mines, in which case flyrock is a 
major consideration. Because of reduced 
burden on the missed hole, violent fly- 
rock may result. Normally, removal of 
the explosive load is recommended in sur- 
face mines. Mucking out a misfire must 
be done with caution, with the minimum 
number of personnel in the area, and 
under the supervision of a competent 
person. 

Removal of stemming and explosives from 
blastholes is more difficult in vertical 
holes than in horizontal holes. Two com- 
mon techniques used are washing material 
out with water or blowing it out with 
air. When the main charge is ammonium 
nitrate and fuel oil (AN-FO), washing 
with water has an advantage because water 



17 



will desensitize the AN-FO. But when re- 
priming of the hole is planned, the use 
of water is a disadvantage and it is bet- 
ter to blow out the stemming with air. 
Also, air has an advantage in that it is 
readily available in many mines. The 
disadvantage of air is that it blows dirt 
and dust into the atmosphere, creating 
poor working conditions around the hole. 
When the explosive is to be removed, both 
air and water may deposit the charge in 
fractures around the borehole, which may 
create more of a problem than if the 
charge is left and dug out during the 
mucking operation. 

The removal of an explosive charge by 
firing a nearby charge is another 



technique that has been used. This meth- 
od is not recommended and should be used 
only as a last resort, because the danger 
of drilling into the misfire is always 
present. In blasts with angled holes, 
such as vee cuts, or where hole length- 
to-burden ratios are high, this technique 
should not be used. When it is used, re- 
covery and disposal of the explosive must 
still be perfornmed in a safe manner. 

If explosives from a misfire must be 
stored, all detonators should be removed 
and stored separately. Explosives and 
detonators removed from a misfire should 
not be stored with other explosives or 
detonators. 



CONCLUSIONS 



Blasting accidents are very costly in 
terms of human suffering, lost produc- 
tion, and damage to equipment. Blasting 
accident can be avoided if all persons 
involved in the blast have a thorough un- 
derstanding of the conditions that can 
result in an accident and an appreciation 
of the hazards involved, and take the 
appropriate precautions. 

The requirements of blast site security 
are common to both surface and under- 
ground mines. The two basic requirements 
of blast guarding are (1) to move mining 
personnel out of range of the blast and 
(2) to prevent the movement of personnel 
back into the blast zone. 

Elements that contribute to the devel- 
opment of safe blast guarding are manage- 
ment commitment to safety, training, def- 
inition of blasting authority, preblast 
planning, definition of blast zone bound- 
aries, selection and placement of guards, 
clearing plans, location of the blast 
initiation site, blasting time, blast 
warning signals, personnel accounting, 
communication of blast location and time, 
and blasting crew communications. 

Many factors can affect the amount of 
flyrock produced by a blast. The blast 
design must be appropriate for the site. 
Blasthole pattern layout and drilling 
must be accurate and must take Into ac- 
count blast-site conditions. All holes 



must be checked before loading of explo- 
sives, and loading must be monitored 
closely so that any problems encountered 
can be corrected. 

Most misfires are due to some problem 
with the initiation system such as fail- 
ure to make a connection, a broken lead, 
or simply not understanding the Initi- 
ation system. Other causes of misfires 
are cutoffs, inadequate priming, and mal- 
functioning of the explosives due to im- 
proper storage. 

Detection of a misfire is no problem if 
none of the holes detonate. However, if 
only a few holes or portions of a single 
hole fail to detonate, detection of the 
misfire can be very difficult. In these 
cases, visual inspection of the muckpile 
for undetonated explosives and boulders 
or other muckpile irregularities that 
suggest possible misfires is the most re- 
liable detection method. 

Disposal of detected misfires is accom- 
plished by removing the explosives with 
water washing or air flushing, repriming, 
and reshooting, or by detonating a nearby 
charge. However, detonating a nearby 
charge can be very dangerous and is not 
recommended. 

The best way to avoid misfire accidents 
and costs is to eliminate their causes. 
This can be done by knowing the charac- 
teristics of the explosives, delays, and 



18 



initiation system, by proper blasting de- 
sign, by taking care in loading the shot 
and hooking up the initiation system, 



and by good housekeeping practices at the 
blasting site. 



REFERENCES 



1. D'Andrea, D. V., and J. Bennett. 
Safeguarding of Blast-Affected Areas. 
Paper in Proceedings of the 10th 
Conference on Explosives and Blasting 
Technique, ed. by C. J. Konya (Lake 
Buena Vista, FL, Jan. 29-Feb. 2, 1984). 
Soc. Explos. Eng. , Montville, OH, 1984, 
pp. 110-119. 

2. Fletcher, L. R. , and D. V. D'Anrea. 
Control of Flyrock in Blasting. Paper in 
Proceedings of the 12th Conference on Ex- 
plosives and Blasting Techniques, ed. by 
C. J. Konya (Atlanta, GA, Feb. 9-14, 
1986). Soc. Explos. Eng., Montville, OH, 

pp. 167-177. 

. A Study of Misfires in Min- 

of the 9th 



1986, 
3. 

ing. 



Paper in Proceedings 



Conference on Explosives and Blasting 
Technique, ed. by C. J. Konya (Dallas, 
TX, Jan. 31-Feb. 4, 1983). Soc. Explos. 
Eng., Montville, OH, 1983, pp. 123-132. 

4. Institute of Makers of Explosives 
Safety Library (Washington, DC). Safety 
Guide for the Prevention of Radio Fre- 
quency Radiation Hazards in the Use of 
Electric Blasting Caps. Pub. 20, Oct. 
1978, 20 pp. 

5. E.I. duPont de Nemours & Co. , Inc. 
(Wilmington, DE). Blaster's Handbook. 
16th ed., 1978, 494 pp. 

6. Dick, R. A., L. R. Fletcher, and 
D. V. D'Andrea. Explosives and Blast- 
ing Procedures Manual. BuMines IC 8925, 
1983, 105 pp. 



19 



BLASTER'S TRAINING MANUAL FOR METAL AND NONMETAL MINERS 
By Michael A. Peltier, 1 Larry R. Fletcher, 2 and Richard A. Dick. 3 



ABSTRACT 



The Bureau of Mines has developed a 
blaster's training manual for the metal 
and nonmetal mining industry. The mate- 
rial is divided into 6 chapters and 47 
modules, with each module covering a 
single topic. (For example, the second 
chapter, which deals with initiation and 
priming, is subdivided into nine modules. 
One module covers initiations systems in 
general, another covers delay series, and 
one discusses priming. The remaining six 
modules deal with each of the six initia- 
tion systems. ) 

The modules were structured to enable 
mine training personnel to easily develop 



a site-specific blasters' training pro- 
gram. Each module contains text material 
that comprehensively covers the topic, as 
well as a paraphrased section highlight- 
ing the major ideas of the text. Also 
included with each module are line draw- 
ings and test questions with answers. 

The objective of this material is to 
increase hazard awareness and foster the 
use of safe blasting practices, with the 
anticipated end result being accident- 
free and productive blasting. 



INTRODUCTION 



Based on accident data obtained from 
the U.S. Mine Safety and Health Adminis- 
tration (MSHA), most blasting accidents 
are caused by human error, lack of hazard 
awareness, or lack of general blasting 
knowledge. A lack of understanding as to 
how explosives function can contribute to 
higher mining costs because of inadequate 
fragmentation or lost production. 

Federal regulations require that every 
person who uses or handles explosive ma- 
terials be experienced and understand the 
hazards involved. Trainees should do 
such work only under the supervision of 
and in the immediate presence of experi- 
enced miners. Federal regulations also 
require hazard and task training for min- 
ers. Most training given on mining 



property is based on experience at that 
mine and is done without the aid of ade- 
quate training materials. An improved 
and more meaningful blasters' training 
program is essential in assisting opera- 
tors to properly train blasters and meet 
MSHA training regulations. 

The blasters' training material was de- 
veloped to aid industry in the prepara- 
tion of a site-specific training course 
and is based on a previous Bureau of 
Mines Information Circular titled "Explo- 
sives and Blasting Procedures Manual". 4 
The intent is to help individuals using 
explosives and blasting agents to develop 
a better understanding of the various as- 
pects of blasting that contribute to a 
safe and efficient blast. 



PREPARING A TRAINING COURSE 



The blaster's training manual has been 
constructed to be easily used in develop- 
ing a site-specific and comprehensive 



Mining engineer. 



''Mining engineer technician. 
3 Staff engineer. 
Twin Cities Research Center, Bureau of 
Mines, Twin Cities, MN. 



blasters course. The material consists 
of discrete modules that contain text ma- 
terial, a paraphrased section, line draw- 
ings, and test questions with answers. 

4 Dick, R. A. , L. R. Fletcher, and 
D. V. D' Andrea. Explosives and Blasting 
Procedures Manual. BuMines IC 8925, 
1983, 105 pp. 



20 



Individual pages have been divided 
lengthwise with the comprehensive text 
material on the left-hand side of the 
page. Each paragraph of the text mate- 
rial is numbered for quick reference. 
The right-hand side of the page consists 
of paraphrased text material with a main 
heading and a paragraph number. The 
person preparing the training course can 
read the paraphrased material quickly in 
order to grasp the main ideas of the text 
material. If an explanation is needed, 
the individual, by noting the paragraph 
number, can go directly to the paragraph 
discussing a particular point. 

Line drawings are included with the ma- 
terial to illustrate specific concepts. 
The line drawings can be easily converted 
to overhead transparencies for use in the 
training course. 

The first step in preparing a blasters 
training course is to determine what ma- 
terial must be covered. This can be ac- 
complished by talking with the blasting 
supervisor and blasters, and by observing 
the blasting operation. To help deter- 
mine what topics need to be covered in 
the course, a checklist is included with 
the material. It is arranged to parallel 
the chapters. By completing the check- 
list, the trainer will be able to lo- 
cate the modules to be included in the 
course. For example, under the "Chapter 



1 — Explosives Products" section of the 
checklist, ammonium nitrate and fuel oil 
(AN-FO) and emulsion maybe noted next to 
the subsection "Blasting Agents." By 
reading the list of modules in the table 
of contents under the "Chapter 1 — Explo- 
sives Products" section, the trainer 
will notice that module 4 discusses AN-FO 
and module 5 discusses emulsions. 

The second step is to gather the train- 
ing material needed. The Information 
gathered from blasting personnel through 
the use of the checklist will indicate 
which modules should be included in the 
course. In addition to the modular mate- 
rial, slides and other visual aids from 
the actual operation should be used. Ad- 
ditional technical information concerning 
specific blasting products can be ob- 
tained from either the explosives sup- 
plier or manufacturer. 

The third step is to write lesson plans 
for the course and arrange the training 
material into a cohesive unit. The writ- 
ing of the lesson plans can be simplified 
by making extensive use of the para- 
phrased sections in the modules. 

Since the experience, knowledge, and 
ability of individual blasters vary 
widely, both the length and amount of ma- 
terial to be included in the course will 
have to be determined by mine 
management. 



CHAPTER CONTENTS 



CHAPTER ONE—EXPLOSIVES PRODUCTS 

Purpose and Description 

The purpose of this chapter is to help 
the blaster develop an understanding of 
various types of explosives. Chemical 
and physical properties of seven types of 
explosive products are discussed. Addi- 
tional information explains nine proper- 
ties of explosives that are used to de- 
termine how an explosive product will 
function under field conditions. Mate- 
rial explaining how to select an explo- 
sive product is included in this 
chapter. 



Objectives 

Upon completion of this chapter, the 
blaster should be able to 

1. Give a concise explanation of the 
nature of various explosive products; 

2. List the basic reactive ingredi- 
ents of an explosive product; 

3. Explain how the detonation pres- 
sure and explosive pressure cause the 
rock to be broken; 

4. Explain the importance of oxygen 
balance as it relates to both the energy 
released and the formation of toxic 
gases; 



21 



5. Describe the individual character- 
istics of the explosive products the 
blaster may be using; 

6. Briefly explain why a particular 
product is being used at the blaster's 
operation; 

7. State and explain nine basic prop- 
erties of explosive products; and 

8. Relate the basic properties of ex- 
plosives to the types of explosive 
products being used on the job. 

Chapter Modules 

Module Title 

1 Chemistry and Physics of 

Explosives. 
2 Types of Explosives and 

Blasting Agents. 
3 Nitroglycerin-Based High 

Explosives. 
4 Ammonium Nitrate-Fuel Oil 

(AN-FO). 
5 Slurries, Water Gels, 

Emulsions. 

6 Heavy AN-FO. 

7 Primers and Boosters. 

8.. Liquid Oxygen Explosives. 

9 Black Powder. 

10 Properties of Explosives. 

11 Explosives Selection 

Criteria. 

CHAPTER TWO— INITIATION AND PRIMING 

Purpose and Description 

The purpose of this chapter is to help 
the blaster develop an understanding of 
six initiation systems. The blaster 
will learn the various components of each 
initiation system, how each individual 
system functions, and the advantages and 
disadvantages of the six systems. Infor- 
mation about the two basic delay series 
and material concerning priming are also 
included in this chapter. 



Objectives 

Upon completion of this 
blaster should be able to 



1. Name the three basic parts of an 
initiation system; 

2. Explain the difference in sensi- 
tivity to initiation between high 
explosives and blasting agents. 

3. State the difference between an 
instantaneous and a delay detonator; 

4. List the various components of the 
initiation system the blaster will be 
using; 

5. Explain how the initiation system 
functions ; 

6. Explain how to check the final 
hookup of the system; 

7. Discuss the potential hazards to 
the initiation system; 

8. Give the definition of a primer; 

9. Name some types of explosives used 
as primers; 

10. Explain the proper procedure for 
making primers; and 

11. Explain why the proper location of 
the primer in the borehole is important. 

Chapter Modules 

Module Title 

12 Initiation Systems. 

13 Delay Series. 

14. Electric Initiation. 

15 Detonating Cord Initiation. 

16 Detaline Initiation System. 

17 Cap-and-Fuse Initiation. 

18 Hercudet Initiation. 

19 Nonel Initiation. 

20 Priming. 

CHAPTER THREE— BLASTHOLE LOADING 

Purpose and Description 

The purpose of this chapter is to exam- 
ine proper blasthole loading techniques. 
The chapter discusses loading procedures 
for both small- and large-diameter 
blastholes. Also included in the chapter 
is material that discusses how to check 
blastholes for proper depth, water, 
voids, and obstructions, and how to miti- 
gate these problems. 



chapter, the 



22 



Objectives 

Upon completion of this chapter, the 
blaster should be able to 



Objectives 

Upon completion of this chapter, the 
blaster should be able to 



1. Explain why blastholes should not 
be loaded and workers should retreat from 
the blast area during the approach or 
progress of an electrical storm; 

2. Describe how to check the borehole 
for proper depth, obsructions, water, and 
voids ; 

3. Explain how to remedy problems, 
such as improper borehole depth, obstruc- 
tions, water, and voids; 

4. State why stemming is important and 
how to estimate the amount of stemming 
needed; 

5. Explain when plastic borehole 
liners or water-resistant cartridges 
should be used; 

6. Explain the proper technique for 
loading the explosive or blasting agent 
the blaster will be using. 

7. Describe the characteristics of the 
type of pneumatic loading the blaster 
will use; 

8. Explain the potential problem of 
static electricity if the blaster is go- 
ing to use a pneumatic loader; and 

9. List the advantages and disadvan- 
tages of using bulk-loaded products in 
large-diameter blastholes. 

Chapter Modules 

Module Title 

21 Introduction. 

22 Checking the Blasthole. 

23 General Loading Procedures. 

24 Small-Diameter Blastholes. 

25 Large-Diameter Blastholes. 

CHAPTER FOUR—BLAST DESIGN 

Purpose and Description 

The purpose of this chapter is to exam- 
ine the factors that influence safe and 
effective blast design. In addition to 
the discussion of design factors for sur- 
face and underground blasting, four con- 
trolled blasting techniques are also 
covered. 



I. Discuss how geology affects 
fragmentation; 

2. Name the most significant geologic 
features to consider when designing a 
blast; 

3. Discuss the importance of a well- 
detailed drilling log; 

4. Explain how to determine the 
burden; 

5. Explain why geologic structure is 
the major factor in determining blasthole 
diameter; 

6. Explain how collar distance affects 
fragmentation size; 

7. Explain the relationship of collar 
distance to airblast and flyrock; 

8. Explain the relationship between 
burden flexing and rock fragmentation; 

9 Discuss the problems of either ex- 
cessive or insufficient subdrilling; 
10. Explain how spacing is determined; 

II. Explain the advantages of milli- 
second delays; 

12. Discuss the two classifications of 
opening cuts; 

13. Explain how to design an angled 
cut; 

14. Explain how to design a parallel- 
hole cut; 

15. Discuss the two types of delays for 
underground blasting; 

16. Name the two main advantages of us- 
ing controlled blasting; 

17. List the four primary methods of 
controlled blasting; and 

18. Discuss the advantages and disad- 
vantages of the various methods of con- 
trolled blasting. 



Module 



Chapter Modules 
Title 



26 Introduction to Blast Design. 

27 Properties and Geology of the 

Rock Mass. 

28 Surface Blasting. 

29. Underground Blasting. 

30 Controlled Blasting Techniques. 



23 



CHAPTER FIVE — ENVIRONMENTAL EFFECTS 
OF BLASTING 

Purpose and Description 

The purpose of this chapter is to 
examine the environmental effects of 
blasting. The material will discuss fly- 
rock, ground vibrations, airblast-, and 
dust and gases. Methods to reduce the 
potential health and safety hazards they 
may present will be discussed. 



CHAPTER SIX—BLASTING SAFETY 

Purpose and Description 

The purpose of this chapter is to help 
the blaster develop a better understand- 
ing of blasting safety, by examining a 
number of auxiliary blasting functions. 
A. number of precautions related to 
previous modules are mentioned. Four ac- 
cident types that occur frequently are 
also discussed. 



Objectives 

Upon completion of this chapter, the 
blaster should be able to 

1. Explain the importance of conduct- 
ing a preblast survey, maintaining 
comprehensive records, and good public 
relations; 

2. Discuss the causes of flyrock; 

3. Discuss methods to alleviate 
flyrock; 

4. Discuss the causes of ground 
vibration; 

5. Discuss design techniques to mini- 
mize vibrations; 

6. State some methods to monitor 
ground vibrations; 

7. Discuss the causes of airblast; 

8. Discuss methods to monitor 
airblast; 

9. List techniques to reduce airblast; 

10. Explain why an adequate amount of 
time must be given for dust and gases to 
be diluted before returning to the blast 
site; and 

11. List the two common toxic gases 
produced by blasting and list techniques 
to reduce them. 

Chapter Modules 

Module Title 

31 Introduction to Environmen- 
tal Effects of Blasting. 

32 Flyrock. 

33 Ground Vibrations. 

34 Airblast. 

35 Dust and Gases. 



Objectives 

Upon completion of this chapter, the 
blaster should be able to 

1. Explain why a knowledge of all cur- 
rent blasting safety regulations is im- 
portant ; 

2. Name the agencies that regulate and 
enforce the use and storage of explosives 
and blasting agents; 

3. Describe the requirements for ve- 
hicles used to transport explosives and 
blasting agents from the magazine to the 
job site; 

4. Explain the importance of marking 
the blast area and keeping nonessential 
personnel away; 

5. Explain when to check for extrane- 
ous electricity; 

6. Discuss why electrical storms are a 
hazard regardless of the type of initia- 
tion system; 

7. Explain the importance of proper 
primer makeup; 

8. List a number of checks to be made 
before borehole loading begins; 

9. Describe various methods to check 
column rise during borehole loading; 

10. Describe some precautions to con- 
sider before and during the hookup of the 
shot; 

11. Explain some good methods for blast 
area security; 

12. Describe the potential hazards to 
check for when reentering the blast site 
after the shot has been fired; 

13. Discuss methods for disposing of 
misfires; and 

14. Discuss the principal causes of 
blasting accidents. 



24 



Chapter Modules 
Module Title 

36 Introduction to Blasting 

Safety. 

37 Explosives Storage. 

38 Transportation From Maga- 
zine to Job Site. 

39 Precautions Before Loading. 

40 Primer Safety. 



Module 



Title 



41 Borehole Loading. 

42 Hooking Up the Shot. 

43 Shot Firing. 

44 Postshot Safety. 

45 Disposing of Misfires. 

46 Disposal of Explosive 

Materials. 

47 Principal Causes of Blast- 
ing Accidents. 



SUMMARY 



A training manual for metal and 
nonmetal raining has been developed by 
the Bureau. This program consists of 47 
modules or topics under 6 major headings 
(chapters). The modules consist of a 
text and outline on a single blasting 
topic, plus questions and answers. 



Supplementing the modules are a 73-itera 
bibliography, a list of regulatory au- 
thorities and their responsibilities, 
additional information on MSHA and OSM 
(U.S. Office of Surface Mining), a glos- 
sary, and 65 illustrations suitable for 
duplication. 



25 



DELAYED BLASTING TESTS TO REDUCE ROCKFALL HAZARDS 
By Virgil J. Stachura 1 and Larry R. Fletcher 2 



ABSTRACT 



The Bureau of Mines conducted delayed 
blasting experiments at a contour coal 
mine, which were designed to reduce over- 
break without special drilling or signif- 
icant additional costs. In the standard 
layout of the blast pattern at this mine, 
the ends of the rows formed the highwall. 
Overbreak was reduced by increasing the 
delays on the last row of holes at the 
highwall, which changed the effective de- 
lay pattern geometry and the direction of 
burden movement. These experiments re- 
sulted in smoother highwalls, which were 
also inherently safer because of the re- 
duced likelihood of rockfall. 

Three delay combinations were tested: 
50 ms longer than the nominal design, 



100 ms longer than nominal, and 50 and 
100 ms longer in the two rows of holes 
nearest to the highwall. The mine's nom- 
inal blast design was a flat V-pattern 
with 17-ms surface delays between holes, 
42-ms surface delays between rows, and 
200-ms in-the-hole delays in each hole. 
All three test designs produced highwall 
improvements, compared with results using 
the nominal design, with occasional ex- 
ceptions because of geologic variations. 
Observations and terrestrial photogram- 
metry showed that the delay changes pro- 
duced generally smoother vertical pro- 
files with less loose material. 



INTRODUCTION 



One of the major hazards found in sur- 
face mining is rockfall from highwalls. 
This hazard occurs in all forms of exca- 
vation in rock, especially where explo- 
sives are used. The explosive energy not 
only fractures the rock to be excavated 
but also damages the rock that borders 
the excavation. This reduces the stabil- 
ity of the highwall and increases the po- 
tential of rockfall. The rockfall hazard 
is normally attributed to blasting prac- 
tices, geologic conditions, and adverse 
weather in 65 pet of accidents resulting 
from fall of rock (1_) . 3 Of these three 
factors, only blasting is controllable, 
and therefore, blasting was the subject 
of this investigation. 

In earlier research sponsored by the 
Bureau of Mines, Engineers International 

' Geophysicist . 

2 Mining engineering technician. 
Twin Cities Research C enter , Bureau of 
Mines, Minneapolis, MN. 

•^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



Inc. evaluated blasting practices at nine 
contour mines (2) • Almost all the mines 
visited had highwall instability prob- 
lems that were aggravated by poor blast- 
ing practices. Engineers International 
conducted eight test blasts that demon- 
strated that good blasting practices did 
improve highwall stability. However, af- 
ter the tests were complete and the con- 
tractor was off the site, the mine per- 
sonnel reverted to their old blasting 
practices. 

To avoid a similar result, the tests in 
this report emphasize simple, easily un- 
derstood changes that minimize economic 
and procedural impact and maximize opera- 
tor acceptance. The experiments are di- 
rected at reducing overbreak without spe- 
cial drilling or significant additional 
cost. They use simple changes In blast- 
hole initiation timing, which improve 
relief by changing the direction and 
time of burden movement. In this report, 
overbreak is defined as excessive break- 
age of rock beyond the desired excavation 
limit (3). 



26 



BLAST DESIGNS 



The approach selected for devising the 
experimental blast designs was to take 
the design in use at the mine site and 
make minor delay-period changes. No oth- 
er parameters were changed intentionally, 
but in the mine's normal blasting proce- 
dures the accuracy of spacings and bur- 
dens varied, more than one hole diameter 
was occasionally used, and powder column 
heights also varied. The mine where the 
tests were conducted used a Nonel 4 (also 
called shock tube) initiating system, so 
the original delay times reflected those 
available using that system. 

The blast design used by the mine, de- 
sign 1, was a flat V with surface delays 
of 17 ms between holes, 42 ms between 
rows, and 200-ms in-the-hole delays (fig. 
1). Designs 2 and 3, experimental de- 
signs used by the Bureau, were the same 
as design 1 except that in-the-hole de- 
lays of 250 and 300 ms , respectively, 



were used in the highwall holes (figs. 
2-3). Design 4, another Bureau design, 
was the same as design 1 except that 300- 
ms in-the-hole delays were used in the 
highwall holes and 250-ms in-the-hole de- 
lays were used in the second row of holes 
from the highwall (fig. 4). Figures 1 
through 4 show the cumulative delay times 
for each blast hole and include arrows to 
indicate the observed direction of burden 
movement. In figures 2 through 4, this 
direction is perpendicular to the plane 
of the highwall, a sign of improved re- 
lief over that illustrated in figure 1. 
Design 4 was tried because of reports of 
overbreak extending far beyond a distance 
equal to one burden ( 4_-5 ) . It was antic- 
ipated that the individual delay in the 
second row of holes from the highwall 
would provide additional relief, reducing 
the damage to the highwall. 



HIGHWALL EVALUATION 



To determine a criterion for evaluating 
the test blasts, discussions were held 
with U.S. Mine Safety and Health Admini- 
stration (MSHA) inspectors, mine super- 
intendents, safety officers, and blast- 
ers. The general theme found in these 
discussions was that a smooth highwall of 
competent appearance is a safer one. A 
competent appearance is achieved by re- 
ducing overbreak. In addition to visual 
inspection criteria, stereo photography 
techniques were used to analyze the 
blasting effects in this study. The use 
of highwall profiles generated from ste- 
reo photographs provided for more analyt- 
ical and consistent comparison of high- 
walls than was possible with visual 
inspection alone. The details of the 
procedure used to take and analyze stereo 
photographs may be found in RI 8916 and 
RI 9008 (.6-2). 

^Reference to specific products does 
not imply endorsement by the Bureau of 
Mines . 



Typical examples of individual profiles 
from two different highwalls are shown in 
figures 5 and 6. Figure. 5 shows a large 
ledge at the top of the test area, a re- 
sult of shot design 1. Figure 6 shows 
an adjacent highwall, which resulted from 
shot design 2 and has a much smoother 
profile. Figures 7 and 8 show a test 
highwall and an array of profiles that 
are the result of shot design 1. Fig- 
ures 9 and 10 show a test highwall and an 
array of profiles that are the result of 
design 2. The photographs (figs. 7, 9) 
were used to evaluate amounts of loose 
material and the appearance of compe- 
tence. The arrays of profiles (figs. 8, 
10) illustrate the degree of overhanging 
material and the size of ledges. 

A total of 59 test blasts and their re- 
sulting highwalls were evaluated by vari- 
ous combinations of stereo photographs, 
on-site observation, photographs taken 
by the blaster, and notes made by the 
blaster. 



27 



. ^*-Next section 
\r of highwoll 



343 326 343 360 377 394 41 




Note Numbers by holes ore 
cumulotive delay times 
Arrow indicates 
burden movement 



FIGURE 1 .—Design 1 : break line prior to 285 ms with 200-ms 
in-the-hole delays throughout. 



343 360 377 394 



rNext section 
of highwoll 



335 352 4I9 




Note: Numbers by holes ore 
cumulative delay times. 
Arrow ind'cates 
burden movement 



FIGURE 2.— Design 2: break line prior to 335 ms with 250-ms 
in-the-hole delays in highwall holes and 200-ms in-the-hole delays 
in remaining holes. 



385 /368 



385 402 4I9 436 



343 326 343 360 377\ 



284 30I 



3I8 335 352 



r 



Next section 
of highwoll 



Note: Numbers by holes are 
cumulative delay times. 
Arrow indicates 
burden movement 



FIGURE 3.— Design 3: break line prior to 385 ms with 300-ms 
in-the-hole delays in highwall holes and 200-ms in-the-hole delays 
in remaining holes. 



427 4I0 427 444 46! 528 595 




Next section 
of highwoll 



Note: Numbers by holes are 
cumulotive delay times. 
Arrow indicates 
burden movement 



FIGURE 4.— Design 4: break line prior to 385 ms with 250- and 
300-ms in-the-hole delays in two rows of holes nearest to highwall 
and 200-ms in-the-hole delays in remaining holes. 



CONCLUSIONS AND RECOMMENDATIONS 



The results demonstrated that greater 
highwall stability can be achieved by 
changing the burden movement to a direc- 
tion closer to perpendicular to the plane 
of the highwall and by allowing more time 
for movement of the burden in front of 
the highwall blastholes. The changes in 
blast design described in this report can 
be implemented without increased costs or 
technical complication. 

The burden movement was redirected by 
using delays in all highwall holes that 



were either 50 or 100 ms longer than 
those previously used in the delay pat- 
tern. Another variation was to increase 
the delay time by 50 ms in the second 
row of holes from the highwall and by 
100 ms in the highwall row of holes. The 
lengthened delays allowed more time for 
the burden to move, thereby reducing 
overbreak, which causes irregular and 
unstable highwalls. At the test sites 
used, the 50-ms-longer delays or a combi- 
nation of 50- and 100-ms longer delays 



28 




FIGURE 5.— Profile from stereo analysis, no additional delays. 
(Numbers indicate distance in feet.) 



FIGURE 6. — Profile from stereo analysis, 50-ms-longer delays 
in highwall holes. (Numbers indicate distance in feet.) 



worked better than when the delay time 
was increased by 100 ms only. This was 
because of the better shearing action 
obtained with the 50-ms incremental in- 
crease. Since the test mine had one par- 
ticular geology and used the Nonel ini- 
tiating system for blasting, other mine 
sites and initiating systems may require 
a different adjustment of the delay time 
to obtain optimum results. The reader 
should also note that, for these tests, 
the highwall was located at the ends of 
the rows rather than the back row of 
holes. 

The test results showed a general im- 
provement in the highwalls even though 
drill-hole alignment and powder column 



heights varied, the geology changed con- 
tinually (because the test site was a 
contour mine), and scaling practices var- 
ied. During the course of study, the 
safety officer and the mine operator both 
observed that the highwalls were notice- 
ably improved and required less cleanup 
time by the dozers used to scale the 
highwalls at the test site. Precise 
evaluation of the blast effects on the 
highwalls proved to be difficult because 
of the variables mentioned above. How- 
ever, after observing the results of 59 
test blasts, the authors recommend an in- 
creased delay of 50 to 100 ms in the 
highwall holes, as described in the 
"Blast Designs" section. 



29 




?-*''*% 



FIGURE 7.— Test highwall, no increase in delay time (blast design 1). 




>Test highwall 



O 5 IO I5 20 

SCALE, ft 



FIGURE 8.— Test highwall profiles, no increase in delay time. 



30 




!&»& 



*tw*ss» 



m 



•i^^m 






FIGURE 9.— Test highwall, 50-ms-longer delays in highwall holes (blast design 2). 




>Test highwall 



. 












20 












15 
10 






\"b 


^ 


Z 


5 


^ i 


\0 




I 


I »x 



5 10 15 20 

SCALE, ft 



FIGURE 10.— Test highwall profiles, 50-ms-longer delays in highwall holes. 



31 



REFERENCES 



1. Theodore Barry and Associates. 
Industrial Engineering Study of Hazards 
Associated With Surface Coal Mines (con- 
tract H0230004). BuMines OFR 48-74, 
1974, 265 pp.; NTIS PB 235 927. 

2. Kendorski, F. S., and M. F. Dunn. 
Safety and Cost Benefits From Improved 
Highwall Blasting Practice (contract 
H0282011, Eng. Inst. Inc.). BuMines OFR 
54-82, 1981, 169 pp.; PB 82-205691. 

3. Dick, R. A., L. R. Fletcher, and 
D. V. D'Andrea. Explosives and Blast- 
ing Procedures Manual. BuMines IC 8925, 
1983, 105 pp. 

4. Hoek, E. Impact of Blasting Damage 
on the Stability of Rock Structures. 
Pres. at 2d Annu. Workshop, Generic 
Miner. Technol. Centr. , Mine Systems 



Design and Ground Control, Reno, NV, Nov. 
12-13, 1984, 10 pp.; available upon re- 
quest from V. J. Stachura, BuMines, 
Minneapolis, MN. 

5. Holmberg, R. , and K. Maki. Case 
Examples of Blasting Damage and Its In- 
fluence on Slope Stability. SveDeFo, 
Stockholm, Sweden, Rep. DS 1981:9, 1981, 
20 pp. 

6. Stachura, V. J., and L. R. Flet- 
cher. Delayed Tests To Improve Highwall 
Stability — A Progress Report. BuMines 
RI 8916, 1984, 24 pp. 

7. Stachura, V. J. , L. R. Fletcher, 
and M. A. Peltier. Delayed Blasting 
Stages To Improve Highwall Stability — 
A Final Report. BuMines RI 9008, 1986, 
12 pp. 



32 



EFFECTS OF BLAST VIBRATION ON CONSTRUCTION MATERIAL 
CRACKING IN RESIDENTIAL STRUCTURES 

By Mark S. Stagg 1 and David E. Siskind 2 



ABSTRACT 



The Bureau of Mines studied the prob- 
lems of blasting-vibration-induced struc- 
tural response and cracking of low-rise 
residential structures in a series of re- 
search projects between 1976 and 1983. 
This paper summarizes the published 
Bureau findings and presents them from 
the point of view of the cracking and 
failure of the construction materials 
used for homes. 

The damage data suggest that, for 
plaster and wallboard attached to the 



superstructure, an increase in the rate 
of cracking is not likely to result from 
blasts generating vibrations of less than 
0.5 in/s. Data on cracks in masonry 
walls suggest that blast-induced vibra- 
tion levels of up to 3.0 in/s may be a 
threshold for local block-length cracks. 
However, additional data are needed to 
quantify vibration level effects neces- 
sary to generate stair-stepped cracks in 
masonry walls, which indicate loss of 
shear load capacity. 



INTRODUCTION 



Ground vibrations from blasting have 
been a continual problem for the raining 
industry, the public living near the rain- 
ing operations, and the regulatory agen- 
cies responsible for setting environmen- 
tal standards. Since 1974, when the 
Bureau of Mines began to reanalyze the 
blast damage problem, several field and 
laboratory studies have been conducted; 
the results of the most recent were pub- 
lished in RI 8969, in 1985 (1_). 3 The 
studies examined blast vibrations with 
respect to generation, propagation, 
structural response, cracking potantial, 
instrumentation, and fatigue (.2~i0* A. 
similar series of studies was conducted 
for airblast (5-6). 



This paper summarizes the material on 
cracking of construction materials used 
in low-rise residential structures; the 
data are excerpted from two comprehensive 
Bureau vibration studies, RI 8507 on dy- 
namic response and damage and RI 8896 on 
fatigue and long-terra influences (_3~4)« 
Specifically, the paper discusses the 
cracking of plaster, wallboard, and mas- 
onry from blasting and other influences, 
giving an overall perspective to the 
blast vibration impacts as part of the 
total lifetime dynamic load for such 
materials. 



CRACKING AND BUILDING PRACTICES 



Current residential construction prac- 
tices address basic human safety and 
not specifically the occurrence of 

' Civil engineer. 

^Supervisory geophysicist. 

Twin Cities Research Center, Bureau 
of Mines, Minneapolis, MN. 

■^Underlined numbers in parentheses re- 
fer to items in the list of references 
at the end of this paper. 



nonstructural minor or cosmetic cracks. 
Many of these practices were derived from 
allowable deflection criteria, in which 
material cracking potential is considered 
(_7-j?). However, cosmetic cracks do de- 
velop, and in 1948, Whittemore (10) dis- 
cussed the lack of guidelines for vibra- 
tions of floors and pointed out that 
"deflection and vibration can be de- 
creased, but only at an increase in 
price." 



33 



LONG-TERM CRACK RATES IN RESIDENTIAL STRUCTURES 



PREVIOUS STUDIES 

Structures crack naturally over time. 
Holmberg (11) analyzed blasting inspec- 
tion reports to estimate a crack rate for 
apartment buildings in Sweden. Two 
apartment buildings were inspected for 
cracks three times between 1968 and 1980. 
The number of observed cracks is plotted 
as a function of time in figure 1. An 
average of 12 to 13 new cracks per year 
occurred for these particular structures. 
Holmberg did not report any specifics on 
the building construction, although con- 
crete is a reasonable assumption. 

The crack rate depends upon the type of 
structure. Rates for 11 wood frame 
houses that were subjected to 26 weeks of 
sonic booms and 13 weeks when there were 
no booms, as reported by Andrews (12) , 
are listed in table 1. Crack rates at 
homes 1 through 4, which were studied 
during both periods, were generally lower 
during the 13-week nonboom period, which 
is similar to Bureau findings discussed 
later. The investigators also found evi- 
dence of the possibility that relative 
humidity and the number of sonic booms 
may together have had an effect on the 
occurrence of cracks. 

The rates of 1.4 to 23 cracks per week 
during the nonboom period are quite high 
compared with the rate observed by Wall 



CKJ\J 


1 1 
KEY 


8 




□ Apartment house 1 






Apartment house 2 




en 
yc 

o 
< 
or 
o 


a 




o 100 




_ 


or 
UJ 

CD 

Z 

n 


o 
o 

D 

1 1 





1965 



1970 



1975 



1980 



FIGURE 1.— Building age versus crack occurrences, after 
Holmberg (7 7). 



(13) in a study of 43 single-story con- 
crete block houses over a 26-week period; 
he reported a crack rate of 2.5 cracks 
per day for the 43 houses (<1 crack per 
week per house). 

The large variation in the crack rates 
reported in the separate studies by Holm- 
berg, Andrews, and Wall is indicative of 
the wide variation of susceptibility of 
houses to cracking. The rates ranged 
from near zero to 23 cracks per week. 
(The yearly rate reported by Holmberg in- 
dicates a cracks-per-week rate of less 
than one.) None of the investigators re- 
ported rates of zero. The large differ- 
ences in the rates reported are partially 
a result of the difficulty of defining 
"cracks". For example, in Wall's report, 
shrinkage cracks were ignored, and only 
new cracks in the moderate (easily dis- 
tinguishable) range were reported. 

These data point out that when months 
pass between preblast and postblast in- 
spections, any postblast inspection is 
likely to find some new cracks that are 
the result of natural aging. 

BUREAU LONG-TERM FATIGUE STUDY 

Blast effects on long-terra crack rates 
were monitored over a 2-yr period at a 
Bureau of Mines test house (4). Bureau 
researchers developed two types of data 
in terms of the expected damage mecha- 
nisms: (1_) fatigue damage from accumu- 
lated exposure, assessed by periodic in- 
spections, and (2_) triggering effects of 
discrete blast events assessed by inspec- 
tions immediately before and after 
blasts, where the strains from blasting 
are added to already existing environ- 
mental strains. Researchers found that 
long-term repetition of the low-level 
blasts (peak particle velocity <0.5 in/s) 
produced no significant effect; however, 
blasts with velocities greater than about 
1.0 in/s were associated with higher 
cracking rates, as shown in table 2. 

The crack rate, or number of new cracks 
per inspection, along with the number of 
blasts that produced ground vibrations 
greater than 0.50 in/s and greater than 
1.0 in/s, is shown in figure 2. Sixty 



34 



TABLE 1. - Crack rates for houses subjected to sonic booms, after Andrews (12). 





Number 














Number o 


f cracks 


House 


of 
stories 


Area, 

ft 2 


Foundation 


Age, 
yr 


Finish 


Occu- 
pied 


per 


week 




Interior 


Exterior 


Boom 


Nonboom 


















period 


period 


1. .. 


1 


1,560 


Concrete 
slab. 


5 


Wallboard. . 


Brick. . . 


Yes.. 


3.7 


1.9 


2... 


2 


1,750 


• • • Q O •••••• 


New 


• • •GO* • • • • • 


... do. . . 


No... 


8.2 


3.3 


3. .. 


1 


1,470 


i * *dOi ♦ • # ♦ # 


8 


• » •QO« • • • • ♦ 


... do* . . 


No. .. 


8.8 


1.5 


4... 


1 


1,160 


Concrete 
stem wall. 


18 


• • •QO» » • • • • 


. . .do. • • 


No... 


6.1 


1.8 


5... 


2 


2,870 


Masonry 
stem wall. 


>50 


Plaster 
and lath. 


Asbestos 
siding. 


No... 


NM 


23 


6... 


1 


1,100 


Concrete 
stem wall. 


25 


i • tGO* • i « • t 


Stone. . . 


Yes.. 


NM 


2.6 


7... 


1 


1,090 


• • • GO »••••• 


30 


Lath and 
wallboard. 


Wood lap 


Yes.. 


NM 


1.4 


8... 


1 


1,280 


» * * QO» * « » » » 


30 


Plaster 
and lath. 


Brick... 


Yes.. 


NM 


3.3 


9... 


2 


2,000 


Masonry 
stem wall. 


40 


Paper on 
plaster 
and lath. 


Wood lap 


Yes.. 


NM 


3.0 


10... 


2 


2,370 


Concrete 
stem wall. 


35 


Plaster 
and lath. 


... do. . . 


Yes.. 


NM 


14 


11... 


1 


1,330 


Concrete 
slab. 


8 


Wallboard. . 


Brick... 


Yes.. 


NM 


2.2 



NM Not measured. 

TABLE 2. - Crack versus vibration (4) 



Blast level, 


in/s 


Cracks per week 




Without corner 


Total 


<0.5 


0.28 
.33 
1.0 


0.84 


>0.5, <1.0 


.89 




1.8 



shots had levels between 0.5 and 
1.0 in/s, while 48 shots had levels above 
1 in/s. Some of the crack rates shown in 
figure 2 include small hairline corner 
cracks, and some do not. The majority of 
corner cracks occurred in the first 8 
months. Cracks were found in nearly 
every corner in the house, but were ig- 
nored until inspection period 15. Then 
it was decided to observe them rigorously 
despite their minuscule size. Corner 
cracks are an inevitable consequence of 
the curing of the tape compound and are 
enhanced by dynamic strains induced by 



human activity. The data that exclude 
corner cracks are more realistic indica- 
tions of blasting influences for homes 
other than new construction, i.e. , within 
6 months. 

Differences were found in the number of 
cracks observed by the two teams of in- 
spectors, Vibration Measurement Engineers 
(VME) and Bureau personnel, during peri- 
ods 1, 15, and 36. The most pronounced 
difference was for period 15. The deci- 
sion to include small corner cracks was 
made after VME had completed its inspec- 
tion for that period but before the Bu- 
reau had completed it inspection for pe- 
riod 15. Other than for that period, 
differences in the number of cracks ob- 
served were an inevitable consequence of 
the difficulty of observing hairline 
(0.01- to 0. 1-mm) cracks. Periods 1, 15, 
and 36 were omitted in calculations of 
crack rates. However, periods in which 
there were unusual external influences, 



35 



CD 

z o 

X 



°o 

<r — ' 

u A 

CD 

2 10 

D I- 

Z O 



20 



15 - 



10 - 



5 - 



A 



■ i i 



^<Nl 



I lfT^.1 I L 



yiVTN iAi i i i/iV. 



' ] i ' ' 




■ Til L 



i I i I i I i I i I i I i I i I i I i 



A 



i i i i 




i i i J I L_l_ 




8 20 22 24 26 28 30 32 34 36 38 40 42 44 46 48 50 

SEMIMONTHLY INSPECTION PERIOD 



FIGURE 2.— Number of cracks and blasts >0.50 in/s and >1.0 in/s versus inspection period. 



including an earthquake and soil removal 
by a scraper 40 ft from the test house, 
were included. The self-triggering seis- 
mograph recorded a 0.06-in/s vibration 
for the scraper activity but did not 
trigger during the earthquake. 

The increase in crack rate with ground 
vibration level indicates that the blast- 
ing produced a triggering strain, at 



about 1.0 in/s. The low crack formation 
rates reported are reasonable since the 
test house was new, showed no differen- 
tial settlement, and was not regularly 
occupied. These conditions resulted in 
low rates of natural crack formation, 
which allowed a few blast-related cracks 
to significantly affect crack formation 
rates. 



CONSTRUCTION MATERIAL CHARACTERISTICS AND CRACKING 



Cosmetic cracks result when a dynamic- 
induced strain (blast vibration or other 
transient vibration) added to a preexist- 
ing strain (static load) exceeds the 
strain level necessary to initiate a 
crack. Differential foundation settle- 
ment, excessive structural loads, and ma- 
terial shrinkage all induce strains that 
can produce random and/or patterned 
cracking. For analyzing blasting ef- 
fects, these strain-inducing forces are 
considered static and the resulting 
strains are called "prestrains. " 

Stress-strain curves are used to de- 
scribe response of materials under load 



up to failure (cracking). Most mate- 
rials, including masonry, plaster, and 
wallboard, respond linearly up to the 
initial yield point. A linear response 
means that deformation (strain) is di- 
rectly proportional to load (stress). 
Beyond initial yeild, plastic deformation 
or creep occurs until ultimate failure 
(fig. 3). The yield point damage is of- 
ten not visually noticeable because of 
limited naked-eye resolution of 0.01 to 
0. 1 mm, particularly in textured surfaces 
such as masonry. 



36 




200 400 600 800 1,000 1,2 00 1,400 1,600 1,800 2,000 2,200 2,400 2,600 2,800 

STRAIN, /xin/in 
FIGURE 3.— Tensile stress-strain curve for 1/2-in-thick wallboard. 



PLASTER 

Plaster was not studied extensively be- 
cause of its widespread replacement by 
wallboard for modern construction. How- 
ever, many of the older homes analyzed 
for RI 8507 (_3) were plastered and pro- 
vided some insight into cracking poten- 
tial. Also, wallboard is a gypsum plas- 
ter faced with paper on both sides. 
Tests run on stripped wallboard are sug- 
gestive of plaster failure (3). Of all 
construction materials, plaster is con- 
sidered most susceptible to damage and 
exhibits fatigue at stress levels less 
than 50 pet of the static failure level 
(14). 

WALLBOARD 

The Bureau studied wallboard cracking 
both in the laboratory and as part of the 
fatigue study of the test house (^0» For 
wallboard in the test house, researchers 
found threshold cracks occurring primar- 
ily in the wall corners and around nail 
heads. They found for wallboard — 

1. The gypsum core failed at strains 
of about 350 yin/in in tension and at 
about 1,000 yin/in in bending, based on 
the nonlinear response points. 

2. For visible cracking, paper failure 
was the controlling factor. Its nonlin- 
ear response point occurred at strains of 
1,000 to 1,200 yin/in (fig. 3). However, 



visual observation of buckling or crack- 
ing was not possible until a slightly 
higher strain level was reached. 

3. Strain rate seemed to affect ulti- 
mate or total failure, but the paper 
yield point was relatively constant. 
This allows comparison of various loading 
factors (e.g. , blasting versus other ac- 
tivities and environmental factors). 

4. Once the wallboard cracked, cyclic 
opening and closing of the crack of up to 
0. 1 mm was observed. These movements 
were unaffected by blasting activities. 

5. Data on cyclic loading behavior of 
wallboard are limited, but results of 
tests on wood products indicated that fa- 
tigue effects can occur at stress (or 
strain) levels equivalent to 50 pet of 
static failure conditions, but over 
100,000 cycles are required. 

MASONRY 

Bureau researchers also studied the 
cracking of concrete block walls, both at 
the test structure with its full-size 
basement and through a series of tests in 
cooperation with the National Bureau of 
Standards (NBS) in Gaithersburg, MD 
(4_, 15). Generally, two types of cracks, 
local and steplike, were identified. Lo- 
cal block-length cracks less than 0.2 mm 
wide were difficult to discern from ex- 
isting mortar joint separations and are 
usually not observed by homeowners. 
Steplike masonry cracks transverse the 



37 



wall along the mortar joint interface 
and, over time, open beyond 0.2 mm in 
width. 

Previous work by Cranston (16) , Green 
(17), and Wroth (1_8) noted that all brick 
walls have small, 0.1-mm cracks upon com- 
pletion. Green stated that 0.1-mm cracks 
are difficult to see and "therefore, do 
not cause concern." As reported by Wood- 
ward (15) , local cracks opened and closed 
throughout the cyclic and monotonic in- 
plane shear tests of a 5- by 5-ft con- 
crete block wall. It was not until a 
steplike crack propagated the length of 
the wall specimen that shear load failure 
occurred. 

Although findings by Bureau researchers 
on masonry failure provide some insight, 
further work at NBS on torsion and out- 
of-plane loading is recommended. Key 
findings for tests with masonry are given 
below. 

1. Observations of tensile cracks at 
strain-monitored sites showed that such 
cracks were first detected visually at 
strain levels well above the first non- 
linear response point because of naked- 
eye limitations (~0.01 to 0.1 mm). 

2. Strains read at the threshold of 
visual cracking using different gauge 
lengths gave different overall strain 
readings as illustrated below. 



but 



0.01 mm 
150 mm 



= 67.0 yin/in, 



Based on the equation e = -= — » 

0.01 mm 
1 3 mm 



Ai. 

= 770 yin/in, 



where 13 and 150 mm are gauge lengths, 
and the visible crack width is 0.01 mm. 
Because strain gauge readings can be mis- 
leading, crack growth is best described 
in terms of displacement. 

3. Local-site strains across the wall 
vary considerably from global strains. 
For in-plane shear failure, global 
strain is measured or calculated across 
the wall diagonally. 

4. Local cracks can occur at low glo- 
bal strains, and global assessment of 
these cracks is not recommended. But, 
for the assessment of steplike cracks 
that propagate across the entire wall, 
the global strain approach appears 
reasonable. 

5. Global failure strain levels for 
steplike cracks are not available. Lim- 
ited testing to date has shown that in- 
plane shear failure may not occur in 
homes because of the relatively light 
vertical load available to prevent rota- 
tion from the shear couple and at least 
a partial conversion of the shear to 
tension. 

6. For cosmetic cracks that do not af- 
fect load-carrying capacity, a crack- 
width criterion has been proposed (17). 
However, the acceptability of crack 
widths varies with material. For con- 
crete, 0.25 mm is the limit of accept- 
ability (19), while 1 mm is the limit of 
acceptability for brickwork (18). 



FACTORS CAUSING STRUCTURE RESPONSE, STRAIN, AND CRACKING 



Bureau researchers studied structure 
responses and cracking associated with 
blasting vibrations in an investigation 
involving a relatively few measurements 
at each of a wide variety of residential- 
type structures (_3). Following this, fa- 
tigue from repeated loading of one house 
over a long period of time was studied 
(4^). For both efforts, measurements were 
made of wall, floor, and racking re- 
sponses, and observations of damage were 
made that could be correlated to specific 
vibration events. A significant part of 
the work was done near large surface coal 



mines with thick soil overburdens and 
large-diameter blastholes, cases which 
had not been studied previously. In all, 
about 900 shots produced useful data on 
structural responses and damage potential 
from blast vibrations. 

ENVIRONMENTAL STRAINS 

Houses are subject to a variety of dy- 
namic loads, in addition to static or 
slightly variable loads from settlement, 
soil changes, and aging. Among the dy- 
namic forces considered significant are 



38 



daily and annual temperature and humidity 
cycles, wind, and human household activ- 
ity. Bureau researchers monitored the 
weather and inside environment during the 
2-yr test period and, in more detail, for 
short periods. For one test, they took 
readings at 3-h increments for a 2-day 
period, simultaneously measuring strain 
at site K.2 , over a major doorway (fig. 
4). Because there were at least four 
factors influencing the strain, re- 
searchers used multiple linear regression 
analyses. Maximum strains from daily en- 
vironmental changes were found to be a 
significant fraction of those needed for 
wallboard core failure or paper cracking. 
The maximum strain observed at K2 was 
+385 uin/in or 39 pet of failure. The 
total maximum strains calculated from the 
correlation equation, assuming the worst 
case for each of the factors, are +675 to 



-817 uin/in, or up to 82 pet of failure. 
"Failure" is defined as the strain level 
of 1,000 uin/in, found to produce wall- 
board cracking as previously discussed 
(fig. 3). 

HUMAN-ACTIVITY-INDUCED STRAINS 

Activities within the home can produce 
significant vibration and strain in local 
structural members (2.~^)' ^ n severe 

cases, such as a hard door slam, the en- 
tire superstructure resonates producing 
strains in every wall, corner, and floor. 
By contrast, nail pounding produces a 
strong response only on the wall af- 
fected. Strains range up to about 100 
uin/in, with typical values being 50 
uin/in in critical areas over windows and 
doorways. 




UJ "■ 

> v- 

•- H 

uj 2 
cr z> 

x 



90 
80 

70 
60 
50 
40 





-— N 


"-.Outside ^*~ 






> 

\ 

s 

X 

• • 






_ — — -" 

s — 




1 


1 1 1 1 1 


1 




1 1 


1 


1 


1 1 



90 
80 
70 
60 
50 
40 



Inside 



' "^^ 4.., ■ 

Outside ■• 



*-c- 



J L 



J L 



J L 



10 
5 


-5 

-10 
-15 

-20 

3(8/10) 



=> uj 
</) a. 
uj in 





""""""• -• •'"" 



JL 



J_ 



24 3(8/11) 6 

TIME.h 



15 



18 



24 



FIGURE 4.— Wallboard joint strain and environmental factors versus time, site K 2 over a doorway. 



39 



300 



270 - 



240 



2 10 



180 - 



■£ 150 

< 
rr 



120 - 



90 



60 



30 



12 3 4 5 6 7 

MAXIMUM GROUND VIBRATION, in/s 

FIGURE 5.— Plaster and wallboard strain versus maximum 
ground vibration at site Si over a doorway. 



- 


I I 


i i 

Envelope_ 


1 1 


- 


- 




line 




- 


- 






Regression / 


- 


- 






^line / 


- 


- 




/ / o 


o 


- 






o / 






- 


/ o 

1 8 ° / 

/ ° o / 

rPX 

/°d° 

05°' 

5%o ! 

i 

, ! , 


o/° o 

/ o 

o 
o o 
o 




- 












1 1 


i i 





BLASTING-INDUCED STRAINS AND 
COMPARISONS 

Blasting responses and strains in resi- 
dential structures were reported in de- 
tail in Bureau RI's 8507 and 8896 C3-4_). 
An example of blast-vibration-induced 
strains from the fatigue study reported 
in RI 8896 is shown in figure 5. Struc- 
ture vibration responses can be transi- 
tional, torsional, vertical uplift, or at 
times a combination of all three. In 
blasting, both the superstructure and 
foundation are typically affected. Non- 
blasting causes of vibration and strain 
act only on the superstructure, except 
for slowly acting soil changes and set- 
tlement. Because initial damage involves 
cosmetic cracks on supersturcture in- 
terior walls, it is appropriate to com- 
pare supersturcture strains from blasting 
and other sources (table 3). These com- 
parisons are only approximate. A given 
vibration level does not always produce 
the same strain even at a single monitor- 
ing point, much less throughout the 
structure, probably because of different 
response modes for different blast 
angles, and wave characteristics. 



TABLE 3. - Comparison of strain levels induced by daily environmental 
changes, household activities, and blasting (4) 



Loading phenomena 


Site 


Induced strain, 
yin/in 


Corresponding 
blast vibration 
level, 1 in/s 


Daily environmental changes 
Do 


Bedroom midwall 
Over doorway. . . 

Over a window. . 

• • • QO » ••••••••• 

• • • GO * ••••••••• 

Over doorway. . . 
Over a window. . 


149 
385 

9 
20 
37 
49 

89 


1.2 
3.0 


Household activities: 


<.03 
.03 




.28 
.50 




.88 



Vibration velocities are based on highest observation strains for a given 
velocity. Use of mean or "typical" values from regression analysis gives ve- 
locities considerably higher . For example, the door slam produces a level of 
strain typically observed at 1.44 in/s ground-measured particle velocity. See 
figure 5. 



40 



OBSERVED CRACKING FROM BLASTING 



As discussed earlier, environmental 
factors induce most of the strain neces- 
sary for the generation of cracks trig- 
gered by household activities or blast- 
ing. Crack rates did not increase until 
blast vibration levels rose above normal 
threshold levels o f 1.0 in/s. It is not 
surprising then that both wallboard and 
plaster cracked at low vibration levels, 
even though failure strain levels for 
wallboard are about 3 times those of 
plaster. 

In reviewing both past and newly avail- 
able data on dynamic vibration response, 
researchers noticed irregular and some- 
times high-amplitude responses when the 
vibration frequencies matched structure 
resonances (fig. 6). A similar effect, 
noticed for the cracking data, was one of 
the most significant findings in RI 8507 
(_3). Consequently, coal mine and quarry 
production blasts that are typically 10 
to 25 Hz produce a greater damage risk 
than smaller scale blasts often used for 
construction, excavation, and secondary 
blasting. 

In a departure from earlier analyses 
and reports, the following review quanti- 
fies damage separately for each of the 
three major construction materials: 
plaster, wallboard, and concrete block. 
The reader is directed to the original 
reports for procedure and analysis de- 
tails O'lll^ 

PLASTER CRACKS 

Threshold and minor cracking data are 
summarized in figure 7 for pre-1975 
studies and in figure 8 for recent Bureau 
research. All these data have been pre- 
viously published in RI 8507 (3) and 
RI 8896 (4^. However, in a departure 
from the earlier reports, these figures 
identify each data point as to source, 
degree of cracking damage, and type of 
material involved. 

Langefors (23) presents the only sig- 
nificant amount of high-frequency data. 
These data, suggest that vibration levels 
as high as 4 in/s may be safe for fre- 
quencies above about 70 Hz. In the de- 
scriptions of damage, in RI 8507's 



table 10 (_3), Langefors did not separate 
the cracking and f all-of-plaster cases. 
Dvorak's study produced observations of 
cracking at some of the lowest peak par- 
ticle velocities, and questions have been 
raised about data reliability. However, 
Dvorak used the same seismic monitoring 
system as Langefors. Dvorak's brick 
structures were likely different from 
Langefors' unspecified structures (proba- 
bly concrete), and the vastly different 
measured frequencies are indicative of a 
soil versus rock foundation. The lowest 
vibration level at which cracking was ob- 
served was 0.5 in/s with Dvorak's data 
and about 0.7 in/s without. 



1 
Corner 


l i i i i 


1 ' 1 


1 1 1 1 1 1 

KEY 




" 






o I -story structure 


- 


- 






A 2- story structure 


- 


. 




O 






- 


A 
A 







- 


- 


A O 


o 


A^ 


- 


i 


A 

Ao A 
A 

O 
i iiii 


o 

O 

° A A O 
. .1 * 


o 
o 

o° A ^ 

° o 9 o°W °° 
Of 8<b$ A o Oj° g 


o - 
o 



Midwoll 




1 ! 




1 


1 I IIII 


1 1 


" 


A 






A 




" 


- 












- 


- 




A 




A 


O o 


- 




A 






A A 
A 


° 








O 




O 
<? Oa 


8 A 

A 












£ 


O 


. 


- 


A 
A 


O 


o 


A 
O 

O 
A 


• '«> : v ° 

o o o 
o«© 2 A 

OB O Q 


- 




O 


i i 


1 


A 

1 


o °o 

o* AO °o t 
A OO 

1 1 IIII 




o 

' 1 



GROUND VIBRATION FREQUENCY, Hz 

FIGURE 6.— Corner and midwall amplification factors (3). 



41 



~i — i — I — I I I u 




mH d do 



KEY 
Ploster cracks and fall of plaster: 
o Dvorak (20) 
• Morrislgl) 
a Thoenen (22) 
o Langefors (23) 
V Jensen (unpublished 
Bureau contract report) 
Wallboord and joint crocks: 
▼ Jensen (unpublished 

Bureau contract report) 
■Wiss (24) 
J ''''I i | | i i_u 



100 



1,000 



FREQUENCY, Hz 



FIGURE 7.— Velocity and frequency levels for threshold plaster 
and wallboard cracking, pre-1975 studies. 



— i — ' ' 1 1| 1 — i — i — i — i i 1 1| 1 

KEY 
o Plaster cracks 
■ Wallboard cracks, all types 
A Wallboard fatigue cracks over nailheads 
V Wallboard fatigue cracks at tape Joints 
* Wallboard panel fatigue 




_j I I I L I I 1 1 



i L-L-UJ 



i i I i i i i 



10 1 00 

FREQUENCY, Hz 



FIGURE 8. — Velocity and frequency levels for threshold plaster 
and wallboard cracking, recent studies (3-4). 



WALLBOARD CRACKS 

The state of cracking of wallboard is 
hard to identify because the interior 
plaster core will crack long before any 
surface effect is visible. Visible 
cracking of paper covering occurs at 



strains about three times those required 
for core failure. Wallboard cracks are 
also influenced by how well panels are 
attached to the superstructure frame. 
Not being structural elements, they are 
not always put under in-plane stress when 
the frame flexes. The core around the 
nailhead is, at best, partially crushed 
upon attachment to the studs, and when 
the studs are uneven major core cracking 
can occur. The response from superstruc- 
ture vibration is additional wallboard 
core crushing around the nailheads, re- 
sulting in a "loose" attachment. 

At the test house, it was observed that 
cracks developed primarily at the plas- 
tered joints at wall corners and in plas- 
ter covering coating over nailheads 
(table 4). The high rate of naturally 
occurring cracks was caused primarily 
from curing of the tape compound. As the 
tests on the structure continued, a de- 
crease of natural frequency of about 
20 pet, e.g., 7.5 to 6 Hz at one loca- 
tion indicated a loss of rigidity and 
general flexure-induced loosening (4_). 

The lowest levels of observed blast- 
vibration-induced cracking occurred at a 
wall corner as crack extensions and when 
a new crack was observed beneath a win- 
dow, at amplitudes of 0.79-1.1 in/s 
(fig. 8). 

Fatigue-induced cracks were observed at 
0.3 to 1.0 in/s. However, this cracking 
required a large number of vibration 
cycles, such as over 50,000 at a 0.5-in/s 
equivalent ground vibration. This 
equates to decades of typical blasting 
with one blast per day producing 10 cy- 
cles per blast. 

MASONRY CRACKS 

Cracks produced in block masonry walls 
by blasting are given in figure 9 for 
past work and figure 10 for recent Bu- 
reau studies (_3~_4)« Most cracks observed 
were local, typically shorter than one 
block length, and about 0.2 mm in width. 
Cracks of this magnitude were observed 
from blast vibrations up to 6.2 in/s and 
were not of concern, being indistinguish- 
able from normal construction and shrink- 
age effects. Their observation is dif- 
ficult, which accounts for the high 



42 



100.0 



10.0 



1.0 - 



I I I I I I I 1 1 



A 
• A 



-I — I — I I 11 III 



i • 



n i i i i 1 1 u 



\ff. 



y 



KEY 

Crack classification: 

A Major damage 
A Minor damage 
• Threshold damage 



J I » i i i I H 



J I i i i i 



J I ' 



10.0 



10 100 

FREQUENCY, Hz 



1,000 



; 


i i i 


1 1 1 M | 


JIM 


- 






- 


. 






_ 


- 






- 


; 




A* 


o 


- 






■ ■ 








/ 


': 




• 


/ KEY 

/ Local masonry wall cracks: 

• Full scale fatigue, Stagg (4) 

a Minor damage-blasting 

■ Model scale fatigue, Kocrncr (27) " 

o Threshold damage -blasting 


/ 


\./ 




• 




i i i 


1 


i i i i i i i il r 



10 100 

FREQUENCY, Hi 



FIGURE 9.— Velocity and frequency levels for local masonry 
wall cracks, pre-1975 studies. 



FIGURE 10.— Velocity and frequency levels for local masonry 
wall cracks, recent studies (3-4). 



TABLE 4. - Wallboard cracks observed in fatigue test house (4). 





Initial 
cracks, 
before 
testing 


Cracks de 


veloped 


Blasting 

level, 

in/s 


Mechanica 


1 shaker tests 1 


Material 


during testing 


Cracks 
developed 


Number of cycles 




Naturally 
occurring^ 


From 
blasting 


at cracking 


Taped corners... 
Taped joints. . . . 


39 
5 
2 
3 


35 
4 
6 
6 


5 
3 

ND 
ND 


0.88-3.5 
1.8 -2.2 

NAp 
NAp 


ND 2 
>3 

1 

1 


NAp 

56,000, 339,500 

56,000 

361,500 



NAp Not applicable. ND None detected. 

Shakers run at resonant frequency at equivalent vibration levels of 0.3 to 1.0 in/s. 

2 Corners almost completely cracked before shaker study. 



TABLE 5. - Masonry wall mortar joint cracks observed in fatigue test house (4) 





Initial 
cracks, 
before 
testing 


Cracks developed 
during testing 


Blasting 

level, 

in/s 


Mechanica 


1 shaker tests 1 


Material 


Number of 
cracks 


Number of cycles 




Naturally 
occurring 


From 
blasting 


at cracking 




20 
21 

NA 
ND 

ND 


28 
11 
NA 
ND 
ND 


7 

1 + 
5+ 
1 
2 


3.4 -6.2 

-6.9 

6.2 -6.9 

.96-1.5 

6.9 


2 
2 

3+ 
ND 2 
1 


229,000, >293,500 


Block 


56,000, >108,500 
>339,500 


Steplike crack. 


NAp 
>339,500 



NA Not available — see text. NAp Not applicable. ND None detected. 
'Shakers run at resonant frequency at equivalent vibration levels of 0.3 to 1.0 in/s. 
2 Existing steplike crack functioned as an area of stress relief. 



43 



number of naturally occurring cracks 
(table 5). Also, these local cracks be- 
came more apparent during the cyclic 
test. Differential motion along the 
block interfaces was easily observed dur- 
ing continued cyclic motion, which ac- 
counts for the low vibration levels, 0.3 
to 1.0 in/s. However, in the test house, 
a blast vibration of 6.9 in/s produced a 
crack of significant magnitude, widening 
a crack beyond the width that was ob- 
served in the absence of a blast. 

SHEAR LOAD FAILURE 

Shear load failure of the basement wall 
of the test house was observed after four 
shots in one day. A diagonal steplike 
crack propagated in the southwest base- 
ment wall, starting at ground level and 
proceeding upward. When these four shots 
were detonated, their vibration levels 
(ranging from 1.0 to 1.5 in/s) were the 
highest recorded in the study up to that 
time. But because observation of cracks 
in masonry is difficult, it remains un- 
known whether blasting or other events 
caused this steplike crack. It is note- 
worthy that no additional steplike crack 



propagations were observed across brick 
or block walls. The existing steplike 
crack functioned as an area of strain re- 
lief during shaker runs. Energy trans- 
mitted by the shakers into the super- 
structure and foundation was primarily 
dissipated in areas of previous cracking. 

Observations were also made of chimney 
and brick veneer responses during cyclic 
shaker tests. The masonry walls were 
relatively stationary, with the super- 
structure cyclically bumping the chimney 
and a brick veneer wall near the roof 
line. Mortar joint cracks developed at 
the chimney-roof interface and horizon- 
tally across the brick veneer just above 
door height. 

Crack data from Edwards and Northwood 
(26) do not specify crack widths. If 
these crack data correspond to observa- 
tions exceeding 0.2 mm (excessive crack 
widths), it would suggest that cracks can 
occur at particle velocity levels of 3 
to 7 in/s with no effect of frequency. 
Additional data are needed to qualify 
frequency effects and the generation of 
stairstep crack patterns across the wall 
signifying shear load failure. 



CONCLUSIONS 



Bureau studies of the response and 
cracking of low-rise residential struc- 
tures from blasting indicated that crack- 
ing of plaster and wallboard is not 
likely below about 0.5 in/s peak particle 
velocity for the worst case of structure 
condition and typical vibration fre- 
quency. 

This safe-level criterion also appears 
independent of the number of blasting 
events and their durations. Researchers 
also noticed that high strains are pro- 
duced in structure walls by normal 
weather conditions, such as wind, temper- 
ature, and humidity cycling. Dynamic 
events such as door slams or blasting 
produce additional strain, which can 
trigger a crack in a structure already 
under strain. Human activities, such as 
door slams, can be equivalent to blast 
vibrations of up to 0.5 in/s. The vibra- 
tion level of 0.5 in/s thus provides a 
minimum value of concern for the impact 



of external transient vibrations on wood- 
frame, low-rise residential structures 
typical of those studied by the Bureau. 

Data on the response and cracking of 
masonry walls from blasting indicated 
that local cracking (block-length) may 
not be noticeable until particle velocity 
levels are up to 3.0 in/s. However, ad- 
ditional research is needed to quantify 
vibration levels that promote the genera- 
tion of stair-stepped cracks that propa- 
gate across the wall and reduce its shear 
load capacity. 

The authors encourage, where possible, 
direct measurements or assessment of 
strains or loads on members likely to 
fail. Alternatively, estimates of re- 
sponses should be based on realistic 
transfer functions relating measured vi- 
brations and reasonably expected re- 
sponses. In particular, applications be- 
yond the scope of the original Bureau 
studies are to be done only with caution. 



44 



REFERENCES 



1. Siskind, D. E., and M. S. Stagg. 
Blast Vibration Measurements Near and On 
Structure Foundations. BuMines RI 8969, 
1985, 20 pp. 

2. Stagg, M. S., and A. J. Engler. 
Measurement of Blast-Induced Ground Vi- 
brations and Seismograph Calibration. 
BuMines RI 8506, 1980, 62 pp. 

3. Siskind, D. E., M. S. Stagg, 
J. W.Kopp, and C. H. Dowding. Structure 
Response and Damage Produced by Ground 
Vibration From Surface Mining Blasting. 
BuMines RI 8507, 1980, 74 pp. 

4. Stagg, M. S., D. E. Siskind, 
M. G. Stevens, and C. H. Dowding. Ef- 
fects of Repeated Blasting on a Wood- 
Frame House. BuMines RI 8896, 1984, 
82 pp. 

5. Siskind, D. E., V. J. Stachura, 
M. S. Stagg, and J. W. Kopp. Structure 
Response amd Damage Produced by Airblast 
From Surface Mining. BuMines RI 8485, 

1980, 111 pp. 

6. Stachura, V. J. , D. E. Siskind, 
and A. J. Engler. Airblast Instrumenta- 
tion and Measurement Techniques for 
Surface Mine Blasting. BuMines RI 8508, 

1981, 53 pp. 

7. Sabnis, G. M. Vibrations of Con- 
crete Structures. Sec. : Introduction 
and Background. Am. Concr. Inst. , 
SP 60-1, 1979, pp. 1-12. 

8. Tuomi, R. I., and W. J* McCutchen. 
Testing of a Full-Scale House Under Simu- 
lated Snowloads and Windloads. U.S. 
For. Serv. , Res. Paper FPL 234, 1974, 32 
pp. 

9. Warwaruk, J. Vibrations of Con- 
crete Structures. Sec. : Deflection 
Requireeraents — History and Background 
Related to Vibrations. Am. Concr. Inst., 
Sp 60-2, 1979, pp. 13-41. 

10. Whitteraore, H. L. , J. B. Cotter, 
A. H. Stang, and V. B. Phelan. Strength 
of Houses — Application of Engineering 
Principles to Structural Design. NBS 
Build. Mater, and Struct. Rep. BMS 109, 
Apr. 1948, 131 pp. 

11. Holmberg, R. , N. Lundberg, and 
G. Rundqvist. Ground Vibration and Dam- 
age Criteria. Constr. Res. Counc , 
Stockholm, Sweden, Rep. R 85:81, 1981, 
30 pp. 



12. Andrews, D. K. , G. W. Zurawalt, 
R. L. Lowery, J. W. Gillespie, and D. R. 
Low. Structure Response to Sonic Booms. 
(U.S. FAA contract FA-64-AC-6-526, An- 
drews Associates Inc. and Hudgins, Thomp- 
son, Ball and Associates Inc. , Oklahoma 
City, OK). Rep. AD618022, Feb. 5, 1965, 
228 pp. ; available from Defense Documen- 
tation Cent. , Alexandria, VA. 

13. Wall, J. R. , Jr. Seismic-Induced 
Architectural Damage to Masonry Struc- 
tures at Mercury, Nevada. Bull. Seisraol. 
Soc. Am., v. 57, No. 5, Oct. 1967, 
pp. 991-1007. 

14. Leigh, B. R. Lifetime Concept of 
Plaster Panels Subjected to Sonic Boom. 
Univ. Tononto, Ontario, Canada, UTIAS-IN- 
191, July 1947, 78 pp. 

15. Woodward, K. A., and F. Rankin. 
Behavior of Concrete Block Masonry Walls 
Subjected to Repeated Cyclic Displace- 
ment. NBSIR 83-2780, 1983, 178 pp.; NTIS 
PB 84-122092. 

16. Cranston, W. B. Masonry Research 
and Codes in the United Kingdom. Paper 
in Earthquake Resistant Masonry Con- 
struction: National Workshop (NBS, Boul- 
der, CO, Sept. 13-16, 1976), ed. by R. A. 
Crist and L. E. Cattaneo. NBS Build. 
Sci. Series 106, Sept. 1977, pp 166-176. 

17. Green, D. G. , I. A. Macleod, and 
W. G. Stark. Observation and Analysis of 
Brick Structures on Soft Clay. Paper 
in Performance of Building Structures 
(Proc. Int. Conf., Glasgow Univ., Mar. 
31-Apr. 1, 1976). Pentech Press, 1976, 
pp. 321-336. 

18. Wroth, C. P. General Report Ses- 
sion IIIA: Response of the Structure 
to Foundation Movements. Paper in Per- 
formance of Building Structures (Proc. 
Int. Conf., Glasgow Univ., Mar. 31- 
Apr. 1, 1976). Pentech Press, 1976, 
pp. 489-508. 

19. Haldane, D. The Importance of 
Cracking in Reinforced Concrete Members. 
Paper in Performance of Building Struc- 
tures (Proc. Int. Conf., Glasgow Univ., 
Mar. 31-Apr. 1, 1976). Pentech Press, 
1976, pp. 99-109. 

20. Dvorak, A. Seismic Effects of 
Blasting on Brick Houses. Pr. Geofys. 



45 



Ustance Cesk. Akad. Ved. No. 169. Geofys. 
Sb., 1962, pp. 189-202. 

21. Morris G. , and R. Westwater. 
Damage to Structures by Ground Vibrations 
Due to Blasting. Mine and Quarry Eng. , 
v. 24, Apr. 1958, pp. 116-118. 

22. Thoenen, J. R. , and S. L. Windes. 
Seismic Effects of Quarry Blasting. 
BuMines B 442, 1942, 83 pp. 

23. Langefors, U. , B. Kihlstrom, and 
H. Westerberg. Ground Vibrations in 
Blasting, Water Power, v. 10, 1958, 
Sept., pp. 335-338; Oct., 390-395; 
Nov., 421-424. 

24. Wiss, J. F., and H. R. Nicholls. 
A Study of Damage to a Residential 
Structure From Blast Vibrations. Res. 
Counc. for Performance of Struct., ASCE, 
New York, 1974, 73 pp. 



25. Stagg, M. S., and D. E. Siskind. 
Repeated Blasting: Fatigue Damaging or 
Not? Paper in Proceedings of the 11th 
Annual Conference on Explosives and 
Blasting Technique, ed. by C. J. Konya 
(San Diego, CA, Jan. 27-Feb. 1, 1985). 
Soc. Explos. Eng. Montville, OH, 1985, 
pp. 96-110. 

26. Edwards, A. T., and T. D. North- 
wood. Experimental Studies of the 
Effects of Blasting on Structures. 
Engineer, v. 210, Sept 30, 1960, 
pp. 538-546. 

27. Koerner, R. M. , and J. L. Rosen- 
farb. Feasibility of Fatigue Assessment 
of Block Walls From Laboratory Scale 
Methods (contract J0285013, Drexel 
Univ.). BuMines OFR 144-80, 1980, 
96 pp.: NTIS PB 81-140139. 



46 



BLAST VIBRATION MEASUREMENTS NEAR STRUCTURES 
By David E. Siskind 1 and Mark S. Stagg 2 



ABSTRACT 



Blasting near structures often involves 
vibration measurement to assess damage 
potential. Several methods of measure- 
ment are used worldwide; however, there 
is no consensus as to which methods are 
technically sufficient and yet practical 
for all situations. 

The Bureau of Mines studied five place- 
ment locations for vibration transducers 
to determine the best method for monitor- 
ing blasting vibrations. The locations 
were — burial in soil next to the struc- 
ture, attachment to the foundation at 
ground level, to the basement floor, or 
to a surface slab, and burial at a dis- 
tance from the structure in undisturbed 
soil. Typical surface mine production 
blasts were used as vibration sources. 



With the exception of the basement 
floor measurements and some of the dis- 
tant measurements, waveforms were similar 
and amplitudes were generally within 10 
to 30 pet of each other. The low- 
frequency part of the wave (5 to 10 Hz) 
was particularly uniform in measurements 
obtained at all five locations. Differ- 
ences in peak values were mostly from 
minor shifts in phase of the high- 
frequency components, which are less sig- 
nificant to structural response and 
potential damage than the low-frequency 
waves. Shallow surface burial resulted 
in good signal detection and the least 
chance of mechanically induced error. 



INTRODUCTION 



The Bureau of Mines studied vibration 
measurement methods applicable to produc- 
tion blasting in surface mines (J_).^ 
Blast vibrations are routinely monitored 
for one of two purposes: (1) to assess 
damage risk for nearby structures and 
(2) to derive predictive equations for 
vibration generation and propagation. 
Despite years of practice and several 
published studies on measurement, the in- 
dustry has not adopted a uniform and con- 
sistent methodology. Occasionally, those 
monitoring blasts fail to obtain repro- 
ducible and accurate vibration records. 

Three measurement methods are in common 
use: (1) direct attachment of the trans- 
ducer to the foundation of the structure 
to be monitored, at or near ground level, 



Supervisory geophysicist. 



o 

•'Civil engineer. 

Twin Cities Research Center, Bureau of 
Mines, Minneapolis, MN. 

-^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



(2) shallow burial of the transducer in 
the soil next to the foundation, and 

(3) measurement on a nearby concrete slab 
such as a driveway or walkway. The 
specific practices followed are often 
based on convenience. However, at suf- 
ficiently low vibration amplitudes, all 
three methods will give similar results. 
The many factors involved, such as trans- 
ducer shape and size, soil or ground 
strength, and density, make for varied 
measurement requirements. Past Bureau 
studies did identify the need to anchor, 
attach, or bury vibration transducers 
for blasts exceeding 0. 2 G acceleration 
amplitude (2_). 

Because of problems of practicality, 
cost, and site access, it is not likely 
that a single measurement method will be 
universally adopted, or needs to be. The 
Bureau, therefore, evaluated the most 
common methods to determine which give 
accurate indications of vibration energy 
transferred into structures while also 
being representative of the blast as a 
vibration source (1). 



47 



Previous research on blast vibration 
monitoring has been concerned with sta- 
bility and slippage (_2~.4)> spiking in 
soil (5-_6)j partial resonances (_7~_8)> im- 
pedance matching (9 ) , Rayleigh wave depth 
effects (6-10), burial in soil, (9, 11- 
12) , and measurement for response spectra 
(13-16). The overall conclusions from 
all these studies are that burial of 
transducers is desirable and that cau- 
tion is required for other methods, 



particularly at high accelerations (com- 
binations of high velocity and high 
frequency). 

The research described in this report 
involves side-by-side comparisons of vi- 
bration time histories from a series of 
surface coal mine production blasts. A 
more complete description of the study 
and related background was published in 
1985 in RI 8969 (1). 



EXPERIMENTAL PROCEDURES 



Four structures were studied at three 
operating mines, two near St. Clairs- 
ville, OH, and one near Evans vi lie, IN. 
All had concrete block or stone basements 
with 5- to 7 -ft depths. Researchers mon- 
itored a total of 23 blasts, containing 
22 to 2,200 lb per delay, at distances 
from 425 to 4,900 ft. 

Monitoring methods varied among the 
sites because of differences of accessi- 
bility and suitability. Concrete slabs 
were near the structures at two sites; 
however, at the other two, the slabs were 
22 and 60 ft away from the main struc- 
ture. Direct comparisons were made as 
closely as possible, by simultaneous mon- 
itoring, with transducers at the follow- 
ing locations: 

1. Buried in the soil with 2 ft of the 
foundation, 

2. Mounted on the foundation wall at 
or near ground level, usually inside. 

3. On the basement floor, 
On a nearby surface slab, a walk- 
driveway, or garage floor, and 
Buried near the structure but re- 

from the influence of disturbed 
around the foundation, 72 to 



4. 
way, 

5. 
moved 
ground 
100 ft distant. 



Special measurements were made of vi- 
bration propagation velocity to examine 
the implications of depth-dependent vi- 
bration amplitudes. Clark had described 
the decrease of Rayleigh wave amplitudes 
with increasing depth (10). For situa- 
tions when the blast vibration is domi- 
nated by the Rayleigh surface waves, 



this could be significant. The decrease 
depends not simply on depth, but on depth 
compared to wavelength (fig. 1). There- 
fore, measurements were made of propaga- 
tion velocity, allowing the calculation 
of wavelengths from 

X = c/f , 

where X = wavelength, ft, 

c = propagation velocity, ft/s, 

and f = frequency, Hz. 

According to Clark, a deeper foundation 
"sees" a decreased amplitude vibration. 
This is also consistent with observed 
low-amplitude vibrations on basement 
floors (10, 16-17). 



-0.5 



RELATIVE VELOCITY AMPLITUDE 
0.5 1 .0 



Horizontal 
component 




Vertical 
component 



1.51- 

DEPTH/WAVELENGTH 

FIGURE 1.— Rayleigh wave amplitude profile versus depth, 
after Clark, Larson and Lande (10). 



48 



RESULTS 



Waveforms were found to be very similar 
for measurements made near and on the 
structure, and also on a nearby surface 
slab. A set of vibration wave compar- 
isons is shown in figure 2. Seven addi- 
tional record sets are shown in RI 8968 
CO. For structure response, the vibra- 
tion's low frequencies (less than 20 Hz) 
are most critical. The low-frequency 
characteristics of the measured records 
were nearly identical, peak for peak and 
wiggle for wiggle. However, high fre- 
quencies present in the outside buried 
gauge records were absent inside, as the 
structure filtered the vibrations by its 
frequency-dependent response. These high 
frequencies provided minor changes in the 
waveform details and peak values. 
Records collected at distances from the 
house differ, mainly from phase changes 
between arriving vibration modes and 
travel-time shifts of energy from various 
parts of the blast. 



Amplitudes varied more than waveform 
frequency characteristics. These ampli- 
tude differences of about 25 pet resulted 
from phase changes and minor amplitude 
differences from the more strongly af- 
fected higher frequencies. In terms of 
structure response, they are not signifi- 
cant. However, they do complicate mon- 
itoring by increasing the scatter in peak 
velocity values. Table 1 summarizes the 
peak vibration amplitudes. As expected, 
basement floors are consistently low, by 
about 25 pet. Most of the peak vibration 
amplitude differences in table 1 are from 
minor phase changes and variations in the 
relatively insignificant high frequen- 
cies. However, they do suggest the ad- 
vantages of using some method of vibra- 
tion analysis to compensate for these 
waveform differences, such as signal 
smoothing, weighting, velocity exposure, 
response spectra, or another frequency- 
dependent method. 



-A/vx/v- 




Location of 
transducer: 

Buried near 
foundation 



Basement wal 



Basement floor 



Midspan of wall 



Time, s 
FIGURE 2.— Vibration records obtained at Fador house, shot 6, longitudinal component of motion, at different locations (1). 



49 



TABLE 1. - Vibration levels measured near and on test structures, peak particle 
velocities, inch per second. 





Shot 


Component 


Measurement location 


Structure 


Ground 


Ground 


Ground 


Basement 


Away from 








level, 


surface, 


level, on 


floor 


structure, 








buried 


on slab 


foundation 




buried 


Schnegg house. . . 


1 


V 


0.32 


NM 


NM 


0.21 


NM 






L 


.26 


0.35 


NM 


.13 


NM 






T 


.28 


.41 


NM 


.15 


NM 




2 


V 


.12 


NM 


0.060 


NM 


NM 






L 


.19 


.14 


.13 


NM 


1 0. 11 






T 


.18 


.12 


.15 


NM 


1 .13 




4 


V 


NM 


NM 


.060 


.054 


NM 






L 


NM 


NM 


.051 


.040 


NM 






T 


NM 


NM 


.079 


.072 


NM 


Do 


5 


V 
L 


.11 
.13 


NM 
NM 


.090 
.14 


.080 
.12 


2 .090 




2 .160 






T 


.11 


NM 


.11 


.080 


2 .160 


Do 


6 


V 
L 


.13 
.18 


NM 
NM 


.12 
.17 


.11 
.15 


NM 




NM 






T 


.13 


NM 


.12 


NM 


NM 


Schnegg house... 


6 


V 


.034 


NM 


.035 


NM 


NM 






L 


.074 


NM 


.075 


NM 


NM 






T 


.070 


NM 


.060 


NM 


NM 


Training house. . 


13 


V 


.017 


.015 


.016 


NM 


3 .020 






L 


.058 


.060 


.035 


NM 


3 .067 






T 


.057 


.059 


.047 


NM 


3 .063 



NM Not measured. 

V Vertical. 

L Longitudinal, 

T Transverse. 



'74 ft f 
2 72 ft f 
3 98 ft f 



rom struc 
rom struc 
rom struc 



ture. 
ture. 
ture. 



CONCLUSIONS 



At the four sites examined in this 
study, the specific measurement methods 
used around structures appear not to be 
critical at low vibration levels. Five 
gauge locations were examined: on a 
surface slab, buried in the ground at the 
structure, mounted on the foundation 
wall, on the basement floor, and buried 
at a distance of 72 to 98 ft from the 
structures. The waveforms for all three 
components, vertical, longitudinal, and 
transverse, were found to be similar for 
the five measurement locations. This was 
particularly true for the low-frequency 
part of the waves, which is of most con- 
cern for vibrational response of struc- 
tures. Low frequency for this study was 
5 to 10 Hz. 



Considerable differences were noted for 
the high-frequency part of the waves, 
mostly in the beginnings, which corre- 
spond to the multiple arrivals from the 
various delayed holes. It is likely that 
these differences were primarily a result 
of the varied wave paths to the different 
monitoring positions, leading to uneven 
and irregular wave interference. The 
high frequencies are of less concern for 
structural response, as discussed in a 
previous Bureau report on structure re- 
sponse from blasting (16). 

Because peak values are influenced by 
the way specific vibration or wave modes 
interact, they were found to vary ir- 
regularly between the different methods. 
However, they were generally within 



50 



20 to 30 pet of each other and not con- 
sistent with the systematic 0.40 factor 
for foundation depths of 6 to 9 ft pre- 
dicted by Clark (10). This demonstrates 
one of the weaknesses of a simple peak- 
particle-velocity criterion and over- 
reliance on interpretation of precise 
values. 

Vibration levels below grade, such as 
on the basement floor, were consis- 
tently lower, suggesting differential 



displacement for the wall top and bottom. 
Apparently, vibrational energy does de- 
crease with depth. 

Rather than recommending a specific 
measuring method, the researchers recom- 
mend that consistency be used at any one 
site. Where the option is available, 
shallow soil burial is still the desired 
method and was found least likely to in- 
troduce additional mechanical error. 



REFERENCES 



1. Siskind, D. E. , and M. S. Stagg. 
Blast Vibration Measurements Near and On 
Structure Foundations. BuMines RI 8969, 
1985, 20 pp. 

2. Stagg, M. S., and A. J. Engler. 
Measurement of Blast-Induced Ground Vi- 
brations and Seismograph Calibration. 
BuMines RI 8506, 1980, 62 pp. 

3. Duvall, W. I. Design Criteria for 
Portable Seismographs. BuMines RI 5708, 

1961, 6 pp. 

4. Fogelson, D. E., and C. F. 
Johnson. Calibration Studies of Three 
Portable Seismographs. BuMines RI 6009, 

1962, 21 pp. 

5. Johnson, C. F. Coupling Small Vi- 
bration Gages to Soil. Earthquake Notes, 
v. 33, Sept. 1962, pp. 40-47. 

6. Gutowski, T. G. , L. E. Witting, 
and C. L. Dym. Some Aspects of the 
Ground Vibration Problem. Noise Control 
Eng. , v. 10, No. 3, 1978, pp. 94-100. 

7. Bycroft, G. N. The Magnification 
Caused by Partial Resonance of the Foun- 
dation of a Ground Vibration Detector. 
Trans. Am. Geophys. Union, v. 38, No. 6, 
1957, pp. 928-931. 

8. Washburn, H. , and H. Wiley. The 
Effect of the Placement of a Seismometer 
on Its Response Characteristics. Geo- 
physics, v. 6, 1941, pp. 116-131. 

9. Duvall, W. I. Design Requirements 
for Instrumentation To Record Vibrations 
Produced by Blasting. BuMines RI 6487, 
1964, 7 pp. 

10. Clark, D. , B. Larson, and G. 
Lande. Vibration: Its Effect and Mea- 
surement Techniques at or Near Buildings. 
Paper in Proceedings of the 9th Confer- 
ence on Explosives and Blasting Technique 
(Dallas, TX, Jan. 31-Feb. 4, 1983). 



Soc. Explos. Eng., Montville, OH, 1983, 
pp. 27-62. 

11. Skipp, B. 0. Ground Vibration In- 
strumentation — A General View. Paper 2 
in Instrumentation for Ground Vibration 
and Earthquakes (Proc. Conf. for Earth- 
quake and Civ. Eng. Dyn. , Keele Univ. , 
Staffordshire, England, July 4, 1977). 
Inst. Civ. Eng., London, 1978, pp. 11-34. 

12. Nicholls, H. R. , C. F. Johnson, 
and W. I. Duvall. Blasting Vibrations 
and Their Effects on Structures. BuMines 
B 656, 1971, 105 pp. 

13. Dowding, C. H. , and P. G. Corser. 
Cracking and Construction Blasting. Im- 
portance of Frequency and Free Response. 
J. Constr. Div. , ASCE, v. 107, No. Co. 1, 
1981, pp. 89-106. 

14. Dowding, C. H. , P. D. Murray, and 
D. K. Atmatzidis. Dynamic Properties 
of Residential Structures Subjected 
to Blasting Vibrations. J. Struct. 
Div., ASCE, v. 107, No. St. 7, 1981, 
pp. 1233-1249. 

15. Langan, R. T. The Use of Single- 
Degree-of -Freedom System as a Dynamic 
Model for Low-Rise Residential Structures 
Subjected to Blast Vibrations. M.S. 
Thesis, Northwestern Univ., Evanston, IL, 
1980, 79 pp. 

16. Siskind, D. E., M. S. Stagg, J. W. 
Kopp, and C. H. Dowding. Structure Re- 
sponse and Damage Produced by Ground 
Vibration From Surface Mine Blasting. 
BuMines RI 8507, 1980, 74 pp. 

17. Siskind, D. E., V. J. Stachura, 
and K. S. Radcliffe. Noise and Vibra- 
tions in Residential Structures From 
Quarry Production Blasting. BuMines 
RI 8168, 1976, 17 pp. 



51 



INITIATION TIMING INFLUENCE ON GROUND VIBRATION AND AIRBLAST 

By John W. Kopp 1 

ABSTRACT 



A major concern with blasting at sur- 
face mines is generation of ground vibra- 
tions and airblast and their effects on 
nearby residences. This Bureau of Mines 
report looks at the use of millisecond 
delays in blast design and their effect 
on the resulting ground vibrations 
and airblast. A total of 52 production 
blasts were instrumented and monitored at 
a surface coal mine in southern Indiana. 
Arrays of seismographs were used to gath- 
er time histories of vibrations and air- 
blast. The data were analyzed for peak 
values of vibration and airblast and for 



frequency content. Various delay inter- 
vals were used within and between rows of 
blastholes. Delay intervals within rows 
were 17 and 42 ms , and those between rows 
ranged from 30 to 100 ms ; these intervals 
are equivalent to burden reliefs of 0.5 
and 1.3 ms/ft within rows and 1.2 to 4.3 
ms/ft between rows. Subsonic delay in- 
tervals within rows reduced airblast by 
6 dB. Large delay intervals between rows 
reduced the amplitude of ground vibra- 
tions; vibration frequency depended pri- 
marily upon the geology of the mine site. 



INTRODUCTION 



Millisecond-delay blasting caps were 
introduced to the mining industry in the 
1940's and gained wide acceptance as a 
tool for improving rock fragmentation. 
The use of these delays also reduced 
ground vibration levels. The Bureau of 
Mines reported on this technique, which 
allowed the explosive in each delay peri- 
od to be treated separately in its con- 
tribution to the ground vibrations, in 
1963 (_1). 2 

The Bureau undertook a major research 
effort during the 1970 ? s to quantify 
these ground vibrations and their effects 
on structures (2). Out of this study 
came the fact that not only amplitude is 
important in preventing damage, but also 
the frequency content of the vibrations, 
because resonances were occurring at the 



natural frequencies of oscillation of 
structures. A study was then undertaken 
at a surface coal mine to determine if 
the predominant frequency of ground vi- 
brations could be controlled by an appro- 
priate choice of blast delay intervals. 
This paper summarizes that study, pub- 
lished as RI 9026 in 1986 (_3). 

During the analysis of the Bureau's 
coal mine data, another study of delay 
control of blasting was undertaken by 
Reil (4^) through a Bureau contract. 
Reil's study in two stone quarries in- 
volved precise timing control. Results 
were mixed with respect to both ampli- 
tudes and frequencies; results also ap- 
peared to be both distance and measuring- 
site specific. 



EXPERIMENTAL DESIGN 



INSTRUMENTATION AND MEASUREMENT 
TECHNIQUES 

Airblast and ground vibrations from 
52 production blasts were measured with 
self -triggered seismographs. These seis- 
mographs recorded three components of 

' Mining engineer Twin Cities Research 
Center, Bureau of Mines, Minneapolis, MN. 



ground motion and the airblast overpres- 
sure on standard cassette audiotapes. 
The tape recorder for each machine was 
automatically activated when the ground 
vibration reached a predetermined level. 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



52 



The frequency range of the transducers 
used to monitor ground vibration was 1 to 
200 Hz. The maximum amplitude that could 
be recorded was 4 in/s. For low-level 
signals, an alternative range could 
be selected with a maximum amplitude of 
1 in/s. 

The airblast channel used a 1-1/8-in 
ceramic microphone. The frequency re- 
sponse of the system was 0.2 to 200 Hz, 
with a maximum peak overpressure of 137 
dB. 

The blasts were also monitored using a 
16-mm high-speed cinecamera. The rotat- 
ing-prism camera was capable of speeds in 
excess of 8,000 frames per second, but a 
rate of 1,000 frames per second was suf- 
ficient for this study. The initiation 
system used was Nonel-* with surface de- 
lays. Nonel tubing was also tied into 
the delay detonators in order to provide 
a flash signal for the camera to record. 
This allowed computation of the firing 
time for each delay to the nearest 
millisecond. 

TEST SITE 

The project test site was a surface 
coal mine in southern Indiana. The mine 
utilized two large draglines to remove 50 
to 100 ft of overburden from a 4- to 5-ft 
coal seam. The overburden was primarily 
shale with some sandstone intermixed, and 
required blasting to facilitate digging 
by the draglines. Blasting was accom- 
plished using 12-1/4-in holes normally 
drilled on a 30-ft-square pattern and 
shot en echelon into a buffer. The ter- 
rain is flat to gently rolling hills. 
The layout of the pit is not influenced 
by topography; it is about 3 miles long 
in a north-south direction. The movement 
of mining is toward the west. 

TEST PROCEDURE 

This series of tests had two objec- 
tives. First, to determine if orienta- 
tion of the shot affected vibration lev- 
els, seismograph arrays were established 

■^Reference to specific products does 
not imply endorsement by the Bureau of 
Mines. 



in four directions from the shot. Each 
array used three instruments, located at 
distances of 300 to 5,000 ft from the 
blast. The complete ground vibration and 
airblast waveforms were recorded at each 
station. From these, the frequency spec- 
tra and peak particle velocities and air- 
blast could be determined. Peak particle 
velocities were plotted versus scaled 
distance for each array direction. A 
least squares fit of the regression line 
was determined for each set of data. A 
one-way analysis of variance test was 
then performed on the data sets to deter- 
mine if the blast parameter under study 
was significant. 

The second objective was to determine 
the effects of varying the delay inter- 
vals between holes and rows. Airblast 
and ground vibration measurements were 
made as before with seismographs deployed 
in arrays in the four directions. Delay 
intervals used were 17 and 42 ms between 
holes in an echelon and 30, 42, 60, 75, 
and 100 ms between echelons. Normal mine 
production shots used 17 ms between holes 
in an echelon and 42 ms between echelons. 
The Nonel Primadet system was used for 
these delays for shots 1 through 36. De- 
lay intervals between rows for shots 37 
to 52 were obtained by using electric 
caps, all of one period, with a sequen- 
tial blasting machine. The high-speed 
camera described previously was used to 
determine actual firing times for each 
hole. Propagation plots were made of the 
airblast and vibration data. 

DIRECTIONAL EFFECTS 

The blast design examined for direc- 
tional effects was the one normally used 
by the mine. A delay of 17 ms was used 
between holes in the echelon, and a delay 
time of 42 ms was used between echelons. 
Ten shots were examined for this experi- 
ment. 4 The ground vibration data are 
shown as a propagation plot in figure 1. 
The airblast data are presented in figure 
2. A least squares regression analysis 

4 Shots 14 through 23. (Shots 1 through 
4 are discussed in the following section; 
shots 5 through 1 3 are not discussed in 
this paper. ) 



53 



10.00 



o 

o 



> 

UJ 

_l 
o 

I- 
ce 

< 

Q_ 



.00 



10 



.01 



10 



± standard 
deviation 




East 



i i i i hi 



10.00 



100 



500 



SQUARE ROOT SCALED DISTANCE, ft/lb'2 



FIGURE 1 .—Propagation plot of peak particle velocity show- 
ing directional differences. 



LOO 



o 

o 



UJ 

> 



cc 
< 
o_ 



.10 



.01 



i 1 r 




South 

a West 

t standard 
deviation 



West 
Sout'h|\^North 

East 



5 10 100 500 

SQUARE ROOT SCALED DISTANCE, ft/lb 2 

FIGURE 3. — Directional effects on propagation of ground vibra- 
tions. All regression lines have a common slope. 



140 



ui 130 
a: 
z> 
</> 

CO 
UJ 

cc 

o_ 

cc 

uj |20 

> 

o 

i- 

co 

< 

_l 

GO 

cc I 10 



100 





KEY 


o 


North 


V 


South 


A 


East 


DDO ° 


West 


* V 


t standard 


\. 


deviation 




J I I I III 



J I I I 'I 



10 100 

CUBE ROOT SCALED DISTANCE, ft/lb 7 3 



1,000 



FIGURE 2.— Propagation plot of peak airblast showing direc- 
tional differences. 



was used to determine the regression line 
of each set of data. Analysis of vari- 
ance tests performed on the vibration 
data determined that the data can be rep- 
resented by four regression lines with a 
common slope (fig. 3). This indicates 
that the vibration level is dependent on 
direction from the blast, but attenuation 
of the vibrations is independent of di- 
rection, with the possible exception of 
the western direction. This may be due 
to a geologic anomaly west of the mine. 
The westen part of the mine is overlain 
by lacustrine and sand and gravel depos- 
its associated with a large creekbed 
drainage area. These deposits tended to 
produce lower predominant frequencies of 
ground vibrations (fig. 4) in che crans- 
verse axis than those produced for the 
other arrays on undisturbed ground (north 
and south directions). Frequencies of 
vertical and radial vibrations did not 
appear to be affected. The frequency of 



54 



vibrations in the reclaimed spoil or 
eastern direction was also predominantly 
1 owe r . 




Eabr station 





West station 




n 1 I I i i I I I 1 i 

5 10 15 20 25 30 35 40 45 

FREQUENCY, Hz 

FIGURE 4.— Predominant frequencies of transverse ground 
vibrations. 



The highest vibration levels were found 
in the western direction, with levels in 
the north array direction the next high- 
est. The results in figure 3 suggest 
that vibration levels in the direction or 
initiation can be twice those in the op- 
posite direction. 

DELAY INTERVALS WITHIN ROWS 

Two different delay intervals were used 
between adjacent holes in each echelon: 
17-ms from shots 1 and 2 and 42-ms delays 
from shots 3 and 4. A 100-ms delay in- 
terval was used between rows. The blasts 
were all shot at the same location in the 
mine using the same blast pattern. 

For these shots, the mine used a square 
pattern drilled on 25-ft centers. The 
pattern was fired en echelon, giving an 
effective burden of 18 ft and spacing of 
35 ft. The actual firing times averaged 
23 and 44 ms for the nominal 17- and 42- 
ms delays, respectively. This gave a 
relief of 0.5 ms/ft of spacing for the 
17-ms-delay shots and 1.3 ms/ft for the 
42-ms-delay shots. The burden delays 
averaged 96 ms , giving a burden relief of 
5.3 ms/ft. 

The direction of the measurement arrays 
from the shot did not appear to signifi- 
cantly affect the airblast data. Analy- 
sis showed the design using 42-ms delays 
produced 6 dB less airblast than the 17- 
ms design (fig. 5). 

An analysis of variance was also per- 
formed for each array direction comparing 
che two delay intervals. The data were 
sufficiently different to require sepa- 
rate regression lines with a common slope 
for the west and north arrays, but showed 
no difference in the east array (figs. 6- 
8). The south array had insufficient 
data for analysis. The direction of ini- 
tiation of the holes in each row was to- 
ward the northwest. The airblast trace 
velocity for the 17-ms delay design was 
supersonic in the north and west direc- 
tions but subsonic in the east direc- 
tion. The airblast trace velocity for 
the 42-ms design was subsonic in all di- 
rections. The airblast from the 17-ms 
design was 7 dB higher in the north array 
and 6 dB higher in the west array, but no 



55 



160 



CE 
Z) 
CO 

to 

LlI 

q: 

0. 

cr 

> 

o 

V- 
co 
< 

ii 

cr 



140 



120 




O I 7ms 
A 42ms 

± standard de 



100 



i i i i i ii 



j i i i ii 



10 



100 



1,000 



CUBE ROOT SCALED DISTANCE, ft/lb ^ 



FIGURE 5.— Propagation plot showing differences in airblast 
levels for the two different blast designs. 




10 100 1,000 

CUBE ROOT SCALED DISTANCE, ft/lb'^ 

FIGURE 6.— Propagation plot of peak airblast for the west 
arrays. 



different in the other directions. This 
would indicate that the reduction in 
airblast is attributable to the fact 
that the trace velocity along the free 
face was subsonic for the longer delay 
interval. 

The two blast designs also show some 
difference in the predominant frequencies 
of the airblast. The design using 17-ms 
delays had more airblast energy in the 
10-Hz range than the 42-ms delay design, 
as shown in figure 9. 

Statistical analysis showed no signifi- 
cant differences in the ground vibration 
levels from the two blast designs. How- 
ever, spectral analysis did show a dif- 
ference in the predominant frequencies 
of the two designs (fig. 10). The 17-ms 
design has its predominant frequencies 
around 10 Hz, while the 42-ms design 
has more scatter in its predominant 
frequencies. 



The delay interval between holes should 
be selected such that the trace velocity 
along the free face is subsonic. This 
resulted in a reduction of airblast of up 
to 6 dB in these tests. The delay inter- 
val between holes did not affect ground 
vibration amplitudes in these tests. 

DELAY INTERVALS BETWEEN ROWS 

Shots 24 through 52 used the same 17-ms 
delay between holes in a row (the average 
value of the actual delay interval be- 
tween holes was 23 ms), but the delays 
between the rows were varied, in five 
steps from 30 to 100 ms. The shot pat- 
tern was 33 ft square, shot en echelon, 
giving an effective burden of 23 ft and 
effective spacing of 47 ft. Five dif- 
ferent delay intervals between rows 
were used to study the effect of burden 
delay timing on vibration levels. The 



56 



160 



160 



Id 

cr 

id 

CO 
CO 

UJ 
cr 

Q. 

rr 

UJ 

> 
o 



co 
< 



CD 

cr 



140 - 



120 





1 1 1 1 1 1 II 

\o 


1 


1 I 1 I 1 1 


- 


OS. 




- 




\A \^ 










A \ 












\ 








A\ 

A N 






KEY 
o I 7 ms 




o\ 

17 




A 42 ms 








- standard deviation 




42 

A 




1 1 1 1 1 1 Ml 


i i 


1 1 i i ii 



100 

10 100 1,000 

CUBE ROOT SCALED DISTANCE, ft/lb 7 3 

FIGURE 7.— Propagation plot of peak airblast for the north 
arrays. 

intervals used were 42 ms , which was the 
delay used by the mine, and 30, 60, 75, 
and 100 ms. The 42 ms was a pyrotechnic 
delay, while the others were selected us- 
ing a sequential blasting machine. Table 
1 gives values of burden relief for the 
different burden delays used. 

Vibration data for each design were 
compared to determine if direction of 
orientation of the seismograph array was 
important. The eastern array (in the 

TABLE 1. - Effective values of burden 
delay intervals 1 



Shot 


Delay interval, ms 


Burden relief, 




Nominal 


Actual 


ms/ft (actual) 


42-44 


30 


27.5 


1.2 


24-36 


42 


48.5 


2.1 


37-41 


60 


58.5 


2.5 


45-49 


75 


76.0 


3.3 


50-52 


100 


99.5 


4.3 



'Effective burden was 23 ft for all 
shots. 



UJ 

rr 
co 

CO 

UJ 

rr 
a 
rr 

UJ 

> 
o 

t- 
co 

< 

_1 

CD 

rr 



140 



120 



100 



~\ — I I III 




KEY 

O 17 ms 
A 42 ms 

i standard dev 



_i i i i ill 



100 
CUBE ROOT SCALED DISTANCE, ft/lb'^ 



1,000 



FIGURE 8. — Propagation plot of peak airblast for the east 
airways. 

spoils) had the lowest vibration levels. 
The highest levels were toward the west, 
where the ground was undisturbed. The 
vibration levels of the other arrays were 
intermediate between these levels. The 
western and northern vibration arrays 
were chosen for further analysis. 

Vibration levels of the different de- 
signs were compared for the north and 
west arrays using regression analysis 
(figs. 11-12). Table 2 gives values of 
intercepts for regression lines with com- 
mon slopes and shows that the two longest 
delay intervals resulted in the lowest 
vibration levels. 

Analysis of the regression lines shows 
that only the two longest delay inter- 
vals resulted in significant reductions 
of ground vibration. This is probably 
the result of reduced confinement be- 
cause sufficient time was allowed by the 
longer delays to move the burden before 
the next echelon of holes was initiated. 
This study found an average vibration 



57 



Q. 



UJ 
CJ 



70 
60 
50 

40 
30 
20 
10 




UJ 

rr 
rr 

=> 60 

o 

o 

° 50 



40 
30 



a 



1 7-ms delay 



in 



i 




QO 



20 
I Oh 



f 



.1 



42-ms delay 







10 15 20 25 

FREQUENCY, Hz 



30 35 



FIGURE 9. — Histograms showing the spectra differences of 
airblast for the two designs. 



o 
Q. 

UJ 
O 

UJ 

rr 
rr 

(J 

o 

o 



1 7-ms delays 




5 10 15 20 25 30 
FREQUENCY, Hz 

FIGURE 10.— Histograms showing the spectra differences of 
ground vibrations for the two designs. 



TABLE 2. - Comparison of regression lines for various burden delay intervals 





Burden delay 
interval, ms 


Array direction 


Regression line 


Regression line 


Shot 


Slope 


Intercept 


with common slope 




Slope 


Intercept 


42-44 


30 
42 
60 
75 

100 
30 
42 
60 
75 

100 




-1.69 
-1.29 
-1.44 
-1.74 
-1.41 
-1.11 
-1.27 
-1.47 
-1.06 
-1.03 


214 
53 

102 

212 
51 
42 
77 

171 
30 
19 


-1.50 
-1.50 
-1.50 
-1.50 
-1.50 
-1.25 
-1.25 
-1.25 
-1.25 
-1.25 


116 


24-36 




104 


37-41 




122 


45-49 




87 


50-52 




71 


42-44 


West 


71 


24-36 




73 


37-41 




80 


45-49 




63 






50 



58 



10.00 



c 



u 

o 



UJ 

> 



UJ 
O 



< 

Q. 




SQUARE ROOT SCALED DISTANCE, ft/lb' 



FIGURE 1 1 .—Propagation plot of peak particle velocity show- 
ing differences caused by burden timing, for the north array. 



>- 

O 

o 



.\J\J 


; 1 




> i 


1 1 1 


i hi i 


i i 


_ 






o 






KEY 




- 










a 30 ms 




- 












□ 42 ms 












D 




60 ms 
v 75 ms 












Sv^j □ 




o 100 ms 










v^ 




D 


± standa 


rd 




.00 


— 








' deviation 


_ = 




- 






\\E D 






_ 




- 






\^o 






- 




" 






D\Wv 






- 




- 








w 9 


^60 
-42 


- 


.10 


^~ 










-30 


-^ 


ns 


1 


ml 


1 


i i 


i i J i 


-75 
-100 

i i 


- 



10 



100 



500 



SQUARE ROOT SCALED DISTANCE, ft/lb ^ 



FIGURE 12.— Propagation plot of peak particle velocity show- 
ing differences caused by burden timing, for the west array. 



reduction of 30 pet using the 100-ms de- 
sign compared with the 42-ms design nor- 
mally used by the mine. 

The frequency content of the ground vi- 
brations was also studied. No reduction 
in low-frequency vibrations was observed, 
which suggests that geology was the 



predominant influence on the frequency of 
vibrations. 

Airblast was also analyzed, but no sig- 
nificant differences were found in fre- 
quency or amplitude as delays between 
rows were varied. 



CONCLUSIONS 



Careful attention to blast design prac- 
tices can help reduce airblast and ground 
vibrations generated by mine blasting. 
This study examined blasthole delay in- 
tervals and their effects on airblast and 
ground vibrations. 

Airblast was influenced by the trace 
velocity along the free face. To reduce 
airblast, the trace velocity, which is a 
function of delay interval and spacing 
between holes in an echelon, should be 
less than the speed of sound in air. 
Airblast was reduced by about 6 dB by 
choosing delays giving a trace velocity 
of 80 pet of the speed of sound rather 
than a supersonic velocity. 

Delays between holes in each row or 
echelon should be greater than 1 ms/ft of 
spacing, in order to prevent reinforcing 



of the airblast wave fronts from the in- 
dividual holes. Care must also be taken 
to avoid selection of delay intervals 
that can cause airblast frequencies equal 
to the natural frequencies of midwalls of 
nearby structures (about 11 to 25 Hz). 
Delay intervals of less than 40 ms will 
usually not present a problem. 

Orientation of the blast and direction 
of initiation had a noticeable effect on 
the magnitude of vibrations. Vibration 
levels in the direction of initiation 
were about twice the level of those away 
from the direction of initiation. Vibra- 
tion levels across the pit from the blast 
were also lower. 

Vibration levels were also dependent on 
the delay interval between rows. Ade- 
quate time must be provided for burden 



59 



relief for each row. This investigation 
found that the delay interval between 
rows should be as long as practical for 
the burden involved. The longest burden 
relief value, 4.3 ms/ft, gave the lowest 
vibration levels. 



The timing of delay intervals between 
rows had no influence on the frequency 
content of the ground vibrations. Geol- 
ogy was the controlling factor for pre- 
dominant frequencies of ground vibrations 
in this investigation. 



REFERENCES 



1. Duvall, W. I., C. F. Johnson, A. V. 
C. Meyer, and J . F. Devine. Vibra- 
tions From Instantaneous and Millisecond- 
Delayed Quarry Blasts. BuMines RI 6151, 
1963, 34 pp. 

2.. Siskind, D. E., M. S. Stagg, J. W. 
Kopp, and C. H. Dowding. Structure Re- 
sponse and Damage Produced by Ground Vi- 
bration From Surface Mine Blasting. Bu- 
Mines RI 8507, 1980, 74 pp. 

3. Kopp, J. W. , and D. E. Siskind. 
Effects of Millisecond-Delay Intervals on 



Vibration and Airblast From Surface Coal 
Mine Blasting. BuMines RI 9026, 1986, 
44 pp. 

4. Reil, J. W. , D. A. Anderson, A. P. 
Ritter, D. A. Clark, S. R. Winzer, and 
A. J. Petro. Geologic Factors Affect- 
ing Vibration From Surface Mine Blast- 
ing (contract H0222009, Vibra-Tech Eng. , 
Inc.). BuMines OFR 33-86, 1985, 204 pp.; 
NTIS PB 86-175858. 



60 



VIBRATIONS FROM BLASTING OVER ABANDONED UNDERGROUND MINES 
By David E. Siskind 1 and Vigil J. Stachura 1 

ABSTRACT 



Vibration wave frequency from surface 
mine blasting is an important influence 
on the response and potential cracking of 
nearby structures. The Bureau of Mines 
studied blasting vibrations in a midwest- 
ern coal mine that occasionally produces 
4-Hz surface waves in its production 
blasting and has received numerous com- 
plaints from neighbors. The mine and 
nearby town are underlain by abandoned 
underground coal mines 85 to 325 ft below 
the surface. 

Blast vibration measurements at the 
site and analysis of mine and regulatory 
agency records indicated that the propa- 
gating medium was primarily responsible 
for the vibration wave characteris- 
tics, including low frequencies, long 



durations, and lower-than-normal attenua- 
tions of amplitude with distance. The 
observed low-frequency waves were consis- 
tent with predictive theoretical models 
of surface wave generation using the 
depths to the old mines. 

Blast designs also contributed to the 
vibrations problem. The complex multide- 
layed blasts generated vibration ampli- 
tudes up to three times those of same- 
weight-per-delay single charges, particu- 
larly for the echelon designs. By con- 
trast, the heavy casting blasts generated 
more of the unusual low frequencies. Be- 
cause of these low-frequency, long-period 
waves, the widely adopted 8-ms minimum 
charge separation criterion may not apply 
at this site. 



INTRODUCTION 



The Bureau of Mines studied a site at 
the western Indiana town of Blanford, 
where surface coal mine blasting was pro- 
ducing unusual low-frequency, long- 
duration vibrations. At the request of 
two regulatory agencies, the Indiana 
Department of Natural Resources (DNR) and 
the U.S. Office of Surface Mining (OSM), 
the Bureau investigated the influence of 
the ground structure, which includes ex- 
tensive abandoned underground workings 
under both the active mining and the town 
of Blanford. These abandoned workings 
are at several depths, the most signifi- 
cant being the extensively mined coalbed 
No. 5 at 225 ft and the partially mined 
No. 4 at 325 ft (fig. 1). The local sed- 
imentary rocks and the coal seams are es- 
sentially horizontal. 

The initial objective of the study was 
to determine if blasting vibrations at 
this site were unusual, as local home- 
owners claimed. Researchers noticed that 

Geophysicist. 
Twin Cities Research Center, Bureau of 
Mines, Minneapolis, MN. 



some vibration records showed data unlike 
any from previous Bureau studies, with 
very low frequencies down to 4 Hz and 
long durations of over 4 s. 

Causes for the unusual data were then 
sought, including the ground and struc- 
ture conditions and factors in the sur- 
face blast designs. By identifying rela- 
tive influences, design guidance could be 
provided to minimize the low frequencies 
generated at this and similar sites. 

Seismologists have recognized that low- 
frequency vibrations are produced in 
soft, weak materials with low propagation 
velocity (J_). 2 Thick soil layers, fill 
areas, and weak sedimentary rocks are ex- 
amples of such materials. Furthermore, 
such layering typically has contrasting 
acoustic properties and produces surface 
waves of several types. These waves are 
inherently low in frequency and develop 
from multiple reflections of conventional 
body waves as they propagate in the rock 

''Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



61 




LEGEND 

Boundary of coalbed 4 mining 
Boundary of coalbed 5 mining 
Roads and streets 
Residence 



1000 



I 


1 
1 




1 


1 
I 




1 ' 


1 


3 


\ 
\ 


1 

1 





Scale, ft 



FIGURE 1.— Town of Blanford, current surface mining highwall, abandoned underground workings, and some of the homes studied 
for vibrations and settlement. 



layers. Ext 
ings also 
layer. 

Recent res 
of delay int 
controlling 
Early resear 
delay effec 
(2) . Howeve 
noticed that 



ensive horizontal mine work- 
provide a strong reflecting 

earch suggests the selection 
ervals as a useful method for 
blast vibration frequencies, 
ch by the Bureau investigated 
ts on vibration amplitudes 
r, these researchers also 
the shorter delays of and 



9 ms (as opposed to 17 and 34 ms) pro- 
duced lower frequencies. 

Research by Anderson (3-4_) and Reil (_5) 
investigated the use of single-charge 
blasts to "calibrate" a site and provide 
a model record for seisraogram syntheses. 
Anderson developed a "Fourmap" scheme, 
which uses superposition of appropriately 
delayed single-charge time histories that 
are then converted into Fourier spectra 



62 



(6). Accurate delay times are assumed. 
This method graphically shows the 
relationship between delays and pre- 
dicted spectral distributions and is 
used commercially by several blasting 
consultants. 

Contrasting with Anderson's and Reil's 
work in rock quarries, a small amount of 
work was done by the Bureau in surface 
coal mines using standard production 
blasts. Here, the thick soil cover and 
relatively weak and complex sedimentary 
rock layers represent a far more influen- 
tial propagation medium. Absorption, 
dispersion, and the generation of second- 
ary waves through multiple reflection 
quickly complicate the wave character and 
make it more difficult to control through 
blast design. These conditions, plus the 
larger holes and charges, tend to gener- 
ate lower frequencies, which are less 
amenable to change. 

Wiss' extensive 1979 coal study did not 
specifically examine delays and vibration 



frequency; however, he did notice the 
occurrence of constructive wave interfer- 
ence when using the 9-ms delay separa- 
tion, but not when using 17-ms delays 
C7_). Kopp specifically sought the influ- 
ence of delay times on frequency for coal 
mine blasts (J3). He concluded that vi- 
bration frequencies did not generally re- 
flect the between-row or within-row delay 
intervals. However, his closest measure- 
ments at this thick overburden site were 
at 300 ft, and it is likely that, at that 
distance and in that material, the pri- 
mary influence on the waves is the propa- 
gation medium itself. In the light of 
current knowledge, Kopp's field program 
in 1981-82 should have included closer-in 
monitoring and tests with more precise 
initiation timing. 

This paper briefly summarizes an in- 
vestigation in the Blanford area reported 
more fully in a recent Bureau Report of 
Investigations (9) and earlier as an un- 
published Bureau report to OSM. 



EXPERIMENTAL PROCEDURES 



Vibration records from 235 production 
blasts were available for analyses of 
amplitudes, frequencies, and total dura- 
tions. These were originally collected 
by the Indiana DNR and the surface mine 
(Peabody Coal Co. , operating the Univer- 
sal Mine) at seven Blanford homes over a 
period of 11 months. Distances of the 
homes from the blast source ranged from 
about 980 to 9,800 ft, and as many as six 
homes were monitored at one time 
(fig. 1). 

For purposes of identifying the genera- 
tion and propagation influences, Bureau 
researchers instrumented seven blasts, 
including two specially fired single- 
charge shots. For these tests, measure- 
ments were made as close as 56 ft to de- 
termine the wave characteristics before 
they were modified by the propagation 
medium. The single charges were used to 
identify the specific vibration- 
generation influences of the complex 
multideck, multidelayed blasts, as con- 
trasted to the simple vibration sources 
represented by the single-charge blasts. 

All vibration records were plotted for 
propagation showing the generation and 



attenuation of particle velocity ampli- 
tudes. Researchers did this for each 
home, for various groups (by neighbor- 
hoods), and as an overall summary for all 
homes. Propagation plots were also pre- 
pared for various blast designs and, in 
the report to OSM, according to vibration 
frequency characteristics. 

In addition to vibration amplitudes, 
frequency characteristics recorded at 
various homes and with various shot de- 
signs were compared. Single charges and 
close-in monitoring revealed the ground's 
natural response frequency of 8 Hz. Shot 
design data were compared for the four 
blasting methods employed by Peabody dur- 
ing the period covered by the records. 
This allowed identification of delay- 
sequencing influences on vibration fre- 
quency, as opposed to ground influence. 

Other site data were collected to iden- 
tify areas around Blanford and the nearby 
surface mine that had been previously un- 
dermined. Depths to these deep coalbeds 
were available from Peabody, which is 
conducting an extensive study of one 
Blanford home. Bureau researchers con- 
ducted level-loop surveys of eight homes 



63 



in September 1985 and the following April 
to identify possible blast-vibration-in- 
duced subsidence and longtime stability. 
Finally, a predictive model was used to 



determine if observed low frequencies 
were consistent with depths to the old 
workings (10). 



RESULTS 



VIBRATION AMPLITUDES 

Propagation plots based on the Peabody 
and DNR reords revealed higher vibration 
amplitudes than were found in previous 
Bureau studies at surface coal mines (_1_). 
Figure 2 is a summary of Peabody- and 
DNR-collected maximum velocities. The 
data are clustered, having been collected 
at seven Blanford homes rather than with 
widely spaced seismograph arrays. Most 
notable is that virtually all the mea- 
surements exceeded the mean of the pre- 
vious coal mine vibration summary (shown 
as maximum velocity, RI 8507, coal). 
Some even exceeded the envelope line of 
highest measurements, which is approxi- 
mately two standard deviations (2a) above 
the mean. 

Separate propagation plots were also 
made for the various blast designs (10). 
Figures 3 and 4 summarize these plots 
based on square root scaled propagation 
and all charge weights within 8-ms 



intervals. Echelon blast amplitudes ex- 
ceeded the historical coal mine summary 



10.00 



1.00 



.10 - 



.01 



= 


i 


1 1 1 Villi 


I 


1 1 


1 1 1 1 l| I 1 1 1 1 


1 11 


- 






Max 


imum 


velocity, 


- 


- 






^RI8507, 


cool 


- 












a 




- 










a 
a 


- 


- 










OA ° ° 


" 


- 












AA AA o 


- 


" 












\ ° °f *£ * 




_ 












\ .'V,' 


. 












\ QCP "O 








KEY 










: 


o 


17 by 100 ms 








- 




A 


17 by 200 ms 






\_ ° °a 














b V° A A 




- 


Q 


17+42 by 200 


ms 




\ ° A 


- 


~ 


V 


17 by 42 ms 






\ 


~ 


- 


' 


i i i i i i il 




i i 


1 1 1 1 l\ 


1 1 1 



10 



SQUARE ROOT SCALED DISTANCE, ft/lb 2 



l/o 



FIGURE 3.— Propagation plot of Peabody and DNR data for 
echelon production blasts. 



10.00 



1.00 



.01 



Mean 




_i i i i i i 1 1 



_i i i i i i i 



10 100 

SQUARE ROOT SCALED DISTANCE, ft/lb^Z 



1,000 



FIGURE 2.— Propagation plot of Peabody and DNR data for all 
homes compared with surface coal mine blasting data summary 
from RI 8507 (1). 



10.00 



1 .00 



.10 



"I — I — I IV 1 I I 



~1 I I I I I I I I 



-Maximum velocity, 
RI8507, coal 



1 1 — I — I — I II u 



_i I | i i i 1 1 




J I i \i i 1 1 



10 100 

SCALED ROOT SCALED DISTANCE, ft/lb' /2 



1,000 



FIGURE 4.— Propagation plot of Peabody and DNR data for 
casting production blasts. 



64 



TABLE 1. - Production blast designs used at the Universal Mine, 
July 1984 through April 1985 



Design type 


Number 
of decks 


Charge weight per delay, lb 


and delays, ms 1 ' 2 


Typical 


Exception or maximum 


Echelon: 

17 by 42 


1 
2-4 
2-4 

1 


1,500 

325 

200- 400 

2,000 


2, 258 maximum. 


17 by 100 


Exception: 625 on 4-22-85 

Exception: 1,475 on 1-05-85 

1,911 on 1-12-85 

3, 842 maximum. 


17 by 200 


Casting: 8, 10, or 
12.5 by 140 to 210.. 



17 by 42 indicates 17-ms delays between holes in a row and 42-ms de- 
lay between rows. 

The 17+42 by 200-ms echelon had too few values to analyze separately. 



from RI 8507 more so than did the casting 
blasts. This appears to be related to 
the blast complexity. Both the casting 
shots and the 17- by 42-ms echelon (17 ms 
between holes in a row and 42 ms between 
rows) were full-column charges. The 
other echelon blasts had 2 to 4 decks 
(table 1). The shots using full-column 
charges produced vibrations close to the 
historical mean, while the other three 
echelon designs produced most of the 
higher amplitude vibrations. 

Although decking is used to reduce the 
charge weight per delay, the increased 
complexity and resulting charge interac- 
tions prevent a corresponding decrease in 
vibration amplitude. Hence, the vibra- 
tion observed from these blasts was 
strong relative to the charge weights per 
delay. These charge weights per delay 
were computed in the traditional manner, 
based on a minimum time separation of 8 
ms. Apparently, this amount of sepa- 
ration is insufficient to separate ef- 
fects for such low-f requency , long-period 
waves. 

It is logical to question if the ab- 
normal amplitudes resulted from the domi- 
nance of low-frequency and inherently 
low-attenuation surface waves instead of 
constructive wave interference between 
delayed charges. Surface waves were 
judged a minor influence on vibration am- 
plitudes (if at all) through the follow- 
ing analysis: 

1. Most of the very low frequency 
cases were from casting blasts. 



2. The casting blast vibration ampli- 
tudes, although high, were closest to the 
historical data of the various blast de- 
signs plotted. 

3. If the high amplitudes were from 
low surface wave attenuation, the cast- 
ing blasts would be the worst case, not 
the least abnormal. Apparently, the 
problem with the casting blasts was not 
their amplitudes but their unusual low 
frequencies and long durations. 

Production and single-charge data were 
obtained on-site from the widely spaced 
array of seismographs (fig. 5). Vibra- 
tion propagations from the single charges 
were close to those of previous Bureau 
studies. However, the production vibra- 
tions, for the same charge weight per de- 
lay, averaged three times higher. Evi- 
dently, vibrations from the delayed 
charges in the multihole, multideck pro- 
duction blasts were interacting. The 
somewhat shallower line slopes for the 
propagation means compared with those 
from previous studies summarized in 
RI 8507, are indicative of a low-vibra- 
tion attenuation with distance. This is 
likely from surface waves being gener- 
ated at the longer propagation distances. 
These results are consistent with the 
previous measurements. Based on con- 
structive wave interference for various 
designs used by Peabody, the number of 
interacting charges were estimated from 
the delay sequencing shown in figures 6 
through 10 (table 2). Sequence times are 
based on nominal delays as shown on the 



65 



30.00 



10.00 - 



o 1.00 - 



.10 



.01 





I I I I I I i 1 1 i ill 


III 


i 


1 1 1 


1 l 1 l 


- 


\^ Production 
v a\/~ Single charge 








- 


; 


\ a \ D 








; 


- 


o \ \ D 








- 


- 


V D \ ° 
\ \ ° 
\ \ a 

\° Dv 








- 


~ 


c\ 


a\ 
□ \ 

o \ 


. a co 






~z 


- 






C \ ° 






- 


- 


KEY 


\ a \ 

e\o D 






- 


- 


o Single-charge shot data 
d Production shot data 




a \ 




- 


: 








D \ 

Q \ 


'- 


- 


i i i i t i i 1 1 i iii 


i i,i 


cA 
1 


i i i 


\ 1 1 1 



10 100 

SQUARE ROOT SCALED DISTANCE, ft/lb'' 2 



1,000 



FIGURE 5. — Propagation plots of recent Bureau tests (9) com- 
pared with surface coal mine summary from Rl 8507 (7). 



tops of figures 6 through 10 and the oc- 
currence of constructive interference 
when delays are within one-half the sinu- 
sodial vibration period. 

FREQUENCY OF VIBRATIONS 

Production Blasts as Vibration 
Source Functions 

All vibration measurements available as 
time history records were analyzed for 
frequency content. Some blasts had very 
low frequencies of about 4 Hz on all 
three components (vertical, longitudinal, 
and transverse) (fig. 6). Other blasts 
had from zero to two components with such 
low frequencies. The source function for 
the blast shown in figure 6 lasted about 
0.7 s, which is also the approximate dur- 
ation of the high-frequency part of the 
vibration record. The low-frequency wave 
began at about 1.7 s and lasted at least 
2 s, for a total vibration duration of 
about 4 s at the measurement distance of 
7,520 ft. 



TABLE 2. - Analysis of charge weights 
for production blasts 

(Maximum number of charges 
per time interval) 



Design type and 

delays, ms 

Echelon: 

17 by 42 , 

17 by 100 , 

17 by 200 , 

Casting: 

Short array i 

Long array , 



8 ms 



17 ms 



60 ms 



7 
9 
7 

5 
13 



Researchers compared vibration rec- 
ords for various blast designs and moni- 
toring locations (figs. 7-10). Although 
the larger casting blasts frequently pro- 
duced low frequencies, the three tested 
echelon designs occasionally also did so. 
Generally, periodicities in the blast 
design, such as between-row spacing, were 
identifiable in some of the records (fig. 
6) but not in others. At the larger dis- 
tances of measurement, the vibrations 
were strongly influenced by the propaga- 
tion medium. Researchers believe that 
low-frequency cases can be moderated by 
minimizing the amount of energy at such 
frquency that is produced by the blast 
design as an energy source (9_)» 

Vibration Wave Types 

The researchers attempted to identify 
the types of waves observed, based on ex- 
pected components and their phases. 
Rayleigh waves are vertically polarized 
with retrograde elliptical particle mo- 
tions. They have significant motion in 
the longitudinal and vertical directions, 
and little in the transverse. The gener- 
ation of these waves requires only a 
single free surface such as the ground- 
air interface. Rayleigh waves can also 
propagate where there is a sharp acoustic 
contracting layer at depth. 

Love waves are horizontally polarized 
shear waves. They are strong only in the 
transverse direction. Generation of Love 
waves requires a layer with top and 



66 



306 


292 


27B 


264 


250 


236 


-o 

222 


208 


510 


496 


482 


46S 


454 


440 


426 


4,2 


724 


710 


696 


662 

o 


668 
— — o 


654 


640 


r 



90S 894 860 866 852 838 824 810 

PATTERN LAYOUT, HERLUDET SYSTEM, TIMES IN MILLISECONDS 



208 ms 

I I I I I I I I 



412 ms 
1 I I I I I I I 



Second row 



SEQUENCE OF CHARGES 



626 ms 
l i 



Third row 



0.1 



Time, s 



509 479 



o ^ — —o o o— 



694 681 667 654 640 627 613 600 586 573 559 546 532 519. 



-O o— 




-o o — — o 



779 765 752 738 725 701 688 674 661 641 634 

PATTERN LAYOUT, HERCUDET SYSTEM, TIME IN MILLISECONDS 



464 ms 

I I I I I I I I I L_l I I I I I I I I I I 



First row 



' ' I ' ' ' ' ' I l i ii 



Second row 



I I I I I I I I I I I I i i Third row 



L_l — I I I I II I li II ii ll li III mill inn i 



^ 



Transverse = 0.051 in/s 



Vertical = 0.078 in/s 



Longitudinal = 0.082 in/s 



VIBRATION RECORDS, HOLLINGSWORTH HOUSE, 7,520 ft 

FIGURE 6.— Casting blasts, 200 ms between rows, 10 ms be- 
tween holes in a row, January 21, 1985. 



rv. 



/V 



^\_ 



SEQUENCE OF CHARGES 
Transverse = 0. 1 I in/s 



Vertical = 0.072 in/s 



Longitudinal = 0.I5 in/s 



I 

Time, s 

VIBRATION RECORDS, MASSA HOUSE, 4,960 ft 

FIGURE 7. — Casting blasts, 200 ms between rows, 10 ms be- 
tween holes in a row, February 16, 1985. 






/42 A84 /I26 /I68 

*59 /lOI /l43 /l85 /227 




'IIS f<l60 ((202 ;<244 

*I77 <4l9 <<26l <^303 ^345^370 

PATTERN LAYOUT, NONEL SYSTEM, TIME IN MILLISECONDS 

First row 



l | Second 


row 


i i i Third 


row 


i i i i Fourth 


row 


l l l i 


i i i i 


i i i i 


1 1 i i 


i ii ii i i i i i i i I Al1 



0.1 



/v. 



/v 



SEQUENCE OF CHARGES 



Transverse = 0.057 in/s 



Vertical =0.018 in/s 



Time, s 



"125 7225 f325 7>25 7525 7625~ 7725 

150 / 250 / 350 /450 / 550 / 650 / 750 / 850 / 950 

175 / 275 / 375 / 475 / 575 / 675 / 775 / 875 / 975 

200 J 300 J 400 J 500 J 600 J 700 / 800 J 900 7 1000 



DECK 
825 ?925 fl,025 DELAY 



1,050 
1.075 
1,100 



T242 7342 7442 7542 7642 7742 7842 7942 71042 

267 / 367 / 467 / 567 / 667 / 767 / 867 / 967 /ip67 

292 / 392 / 492 / 592 / 692 / 792 / 892 / 992 / 1092 

317 i 4,17,/ 517 A 617 ,/ 717 j, 817,/ 917,/ 1,017 ± 1,117 

259 359 459 559 659 759 859 959 1059 

284 384 484 584 684 784 884 984 1,084 

309 409 509 609 709 809 909 ip09 I'l 09 

334 434 534 634 734 834 934 1,034 ft 34 

PATTERN LAYOUT, NONEL SYSTEM, TIME IN MILLISECONDS 



No 5 
No 6 
No 7 
No 8 



125 ms 
I I I 



Top detk 



I Second deck 



J I L 



— First row »| 

i i i i i i 1 1 i i I I I I I 1 1 I I I I I i i i I I I i : i I I I i i i I i i M I I i i i i All 



0.1 



SEQUENCE OF CHARGES 



Time, s 



Transverse = 0.39 in/s 

-vvVv^AAAAAAArv^^^- — 



Vertical - 0.22 in/s 



Longitudinal = 0.35 in/s 



Time, s 
VIBRATION RECORDS, JACKSON HOUSE, AT 8,510 ft 

FIGURE 8.— Echelon blast, 42 ms between rows, 17 ms be- 
tween holes in a row, January 4, 1985. 



Time, s 
VIBRATION RECORDS, JACKSON HOUSE, AT 1,320 ft 

FIGURE 9.— Echelon blast, 100 ms between rows, 17 ms be- 
tween holes in a row, March 4, 1985. 



"275 r483 /687 7"89l A095 / 1,299 

303 /Sit /7I5 /9I9 / 1.123 / 1.327 

210 331 / 539 / 743 / 947 / I.I SI / 1.355 




DECK 

DELAY 



/505 /709 /9I3 /I.II7 / 1,321 

533 / 737 /94I / 1,145 / 1,349 

561 / 765 / 969 / 1,173 / 1,377 

527 "731 "935 1,139 "1,343 

555 759 963 1,167 1,371 

583 787 991 1,195 1,399 



4,459 



PATTERN LAYOUT, HERCUDET SYSTEM, TIME IN MILLISECONDS K 



No 6 



No 8 



1 1 III 


1 i i Top deck 


i i ill III Second deck 


I , III II i Bottom deck 


ill III 


1 II III II 1 *" 



First row Second row 
(one hole) (one hole) 



Fourth row 




SEQUENCE OF CHARGES 



Tronsverse c 0.072 in/s 



Verticol =0.033 in/s 



Longitudinal =0.053 in/s 



65ft 



67 



3.0 (in/s)/in 




226ft-^ 



407ftJj|4j\A 



1 .0 (in/s)/in 



0.5 (in/s)/in 



807 



ft^'iHvVVW- 



0.4 (in/s)/in 



VIBRATION RECORDS, JACKSON HOUSE, AT 7,990 ft 

FIGURE 10.— Echelon blast, 200 ms between rows, 17 ms be- 
tween holes in a row, January 12, 1985. 



I,I65 ft — ■ JVV- 



0.3 (in/s)/in 



bottom boundaries having good reflect- 
ing properties. Extensive underground 
voids could provide such a reflecting 
surface, as could any low-velocity layer. 
The amounts of low frequency appeared to 
be related to the distance from the 
blast; however, trends were not consis- 
tent from site to site. 

Site Differences for Vibration Waves 

Six of the seven measurement sites 
had patterns of low frequencies resem- 
bling Rayleigh waves or Rayleigh and 
Love waves together. Two of these sites 
also occasionally had only high frequen- 
cies (e.g., 20 Hz) at the same time and 
for the same shots when low frequencies 
were being measured elsewhere. One site 
(Jackson) had every kind of combination 
at different times. The seventh site 
(Verhonik) had only transverse low fre- 
quencies, suggesting Love waves. This 
seventh site was in a different location 
from the others, east of the blasting 
rather than north. There was a 100- 
ft-thick layer of mine spoils between 



■'■A/z/uVn/V 



0.1 (in/s)/in 



2515 ft - 



FIGURE 1 1 .—Vibration records from single charge, vertical. 
Horizontal scale is 500 ms/in. 



the site and 
marizes all 
records. 



the blasting. Table 3 sum- 
available time history 



Single-Charge Blasts 

The two special single-charge shots 
were fired to identify the blast initi- 
ation-sequencing influences on the wave 
forms as well as the previously discussed 
vibration amplitudes. They also showed 
the complexity of the propagating medium. 
Figure 11 is the vertical vibration rec- 
ord from a 1-m column of explosive that 
took about 0.3 ms to fully detonate. Af- 
ter propagating 65 ft, the vibration 
duration was over 150 ms. At a distance 
of 1,165 ft, the vibration duration was 
dominated by a 6.5-Hz wave lasting about 
2 s. This observation suggests that any 
blast at this site is a potential 



68 



TABLE 3. - Vibration components showing very low frequencies (~4 Hz), 1 
by measurement site and blasting method 



Date 



Residence 



Massa 



Volk 



Hollings- 
worth 



Zell 



Polomski 



Jackson 



Verhonik 



ECHELON BLASTS, 17 ms BETWEEN HOLES IN A ROW, 42 ms BETWEEN ROWS 



12-18-84. 
12-19-84, 
12-22-84. 
12-28-84. 

1- 4-85. 
1- 9-85. 



HF 
L.V 

V 

V 
111 



HF 
HF 

HF 
HF 



ECHELON BLASTS, 17 ms BETWEEN HOLES IN A ROW, 100 ms BETWEEN ROWS 



3- 2-85. 
3- 2-85. 
3- 2-85. 
3-14-85. 

3-16-85. 
3-18-85. 
3-21-85, 
3-25-85, 

3-26-85, 
9-10-86. 
9-10-86. 
9-10-86. 

9-12-86. 
9-13-86. 



HF 



HF 

8 Hz 

HF 



HF 



HF 



HF 

HF 

-5 Hz 

HF 



HF 

HF 

HF 

8 Hz 

8 Hz 
HF 
HF 
HF 

HF 
•10 Hz 
'10 Hz 
■14 Hz 

~9 Hz 



ECHELON BLASTS, 17 ms BETWEEN HOLES IN A ROW, 200 ms BETWEEN ROWS 



7- 2-84. 
7- 3-84. 
7- 6-84. 
7- 7-84. 

7- 7-84. 
7-11-84. 
7-11-84. 
7-11-84. 

7-13-84. 
7-14-84. 
7-28-84. 
1- 5-85. 

1-12-85. 
3- 5-85, 
3- 6-85. 
3- 7-85, 

3- 9-85. 



HF 



HF 



L,V 



L,V,T 



T 

T 

L,V,T 

L,V,T 

L,V,T 
T 
T 
T 

T 

T 

L,V,T 



HF 
HF 
HF 

HF 



See explanatory notes at end of table. 



69 



TABLE 3. - Vibration components showing very low frequencies (~4 Hz), 1 
by measurement site and blasting method — Continued 





Residence 


Date 


Mass a 


Volk 


Hollings- 
worth 


Zell 


Poloraski 


Jackson 


Verhonik 



CASTING BLASTS 



8- 9-84. 

8-13-84. 
20-30-84. 
11- 5-84. 

11- 7-84. 
11-10-84. 
11-15-84. 
11-17-84. 

11-19-84. 
11-23-84. 
11-24-84. 
11-26-84. 

12- 1-84. 
12- 4-84. 
12- 6-84. 
12- 8-84. 

12-10-84. 
12-12-84. 
12-15-84. 
12-17-84. 

1-12-85. 
1-14-85. 
1-17-85. 
1-21-85. 

1-25-85. 
1-28-85. 
1-31-85. 
2- 2-85. 

2- 6-85. 
2- 9-85. 
2-14-85. 
2-16-85. 



2-19-85, 
2-21-85, 
3- 1-85. 



L,V,T 




L,V,T 




L,V,T 




L,V,T 






L,V,T 


L,V,T 


L,V,T 




L,V,T 


L,V,T 


L,V 


L,V,T 


T 




L,V,T 




V 




L,V,T 




L,V 


L,V,T 


L,V 




L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V,T 


L,V 


L,V,T 


L,V 





L,V,T 
L,V,T 

L,V,T 
L,V,T 

L,V,T 



L,V,T 
L,V,T 
L,V,T 

L,V,T 

L,V,T 

L,V 

L,V,T 

L,V,T 
V 



L,V,T 
L,V,T 



L,V,T 

L,V,T 
L,V,T 
L,V,T 



L,V,T 
L,V 



V 
V 

L,V 
L,V,T 

L,V 

L,V,T 
L,V 
L,V 
L,V 



L,V 



V 

V 

L,V 



L,V 
HF 



T 
T 

T 2 
T 2 
T 2 



T 
T 
T 
T 

T 
T 
T 
T 

T 

T 
T 2 

T 2 



r L Longitudinal (or radial). V Vertical. T Transverse. 
Rayleigh surface waves would dominate L and V components; Love surface waves would 
dominate T component. 

HF Higher frequency vibration with no clear components below 10 Hz. 
Vertical record may be defective; may have been present, but was not visible. 



NOTE. — No entry means no records available. 



70 



TABLE 4. - Summary of two level loop surveys of eight 
Blanford residence. 



House 




Maximum elevation 
differences, ' ft 


Maximum angular 
distortion 2 




Sept. 1985 


April 1986 


Sept. 1985 


April 1986 


Ahlmeyers. . 




0.06 
.12 
.08 
.09 
.30 
.09 
.06 
.37 


0.07 
.11 
.07 
.11 
.26 
.09 
.05 
.39 


1/340 
1/250 
1/300 
1/200 
1/108 
1/560 
1/250 
1/65 


1/340 
1/260 
1/340 
1/164 
1/125 
1/533 
1/300 
1/65 


Finger, E.. 
Finger, 0.. 
Jovanovich. . 





1 Accuracy is ±0.01 ft. 

2 l/340 = Distortion of 1 part in 340. 



low-frequency problem as the ground there 
favors such frequencies. 

Theoretical Model 

The O'Brien model computes surface 
waves generated by multiple reflections 
of compressive body waves in a low-veloc- 
ity layer (10). Another version by Gupta 
is for shear waves. For a strong veloc- 
ity contrast (strong reflector), the sim- 
plified relationship is 



T = 



4h 



where T is the surface wave period, h is 
the layer thickness, and V] is the propa- 
gation velocity for the low-velocity lay- 
er. Assumed for this simple model is 
that V, << V 2 , with V 2 being the high- 
velocity layer. 

At Blanford, researchers measured a V 2 
of 10,000 ft/s and, using a 1.6-s arri- 
val time difference, calculated a Vj of 
2,700 ft/s. For a 6.5-Hz surface wave 
(T = 0.15 s), an h of about 102 ft is in- 
dicated. Similarly, for a 4-Hz surface 



wave, h is 167 ft. Depths to the coal- 
beds as measured by Peabody at one site 
ranged from 85 to 394 ft. The exten- 
sively mined No. 5 was at 226 ft. 

LEVEL-LOOP SURVEYS OF HOMES 

Researchers surveyed eight Blanford 
home foundations to identify surface and 
structural changes from possible subsid- 
ence over abandoned workings (9). Only 
one of the eight homes was also monitored 
for vibration (Zell house). The primary 
selection criterion was the existence of 
a clear foundation horizon for surveying. 

Most of the homes were out of level by 
significant amounts of up to one part in 
65 (table 4). Boscardin cites that angu- 
lar distortions or deflection ratios of 1 
part in 300 can cause cracking in panel 
and load-bearing walls (11). A resurvey 
7 months later showed no appreciable ele- 
vation changes. With these data alone, 
it is not possible to tell if the homes 
are distorted or strained, or if these 
differences were originally built in. 
Apparently, no significant long-term 
changes are occurring. 



CONCLUSIONS 



The propagating medium appears pri- 
marily responsible for the adverse vibra- 
tion impacts in Blanford, through three 
mechanisms: (1) It favors generation of 
low-frequency surface waves of several 
types, with frequencies between 4 and 10 
Hz, (2) it has the appearance of reduced 



vibration attenuation (higher amplitudes) 
with distance compared with other coal 
mine sites, and (3) it produces interac- 
tions between delayed charges beyond 
those expected from the blasts as de- 
signed, because of constructive wave in- 
terference for these long-period waves. 



71 



Although further study of the subsur- 
face conditions is needed in order to 
completely understand all the factors, 
the observed surface waves are consistent 
with a strongly reflecting subsurface in- 
terface at a depth of about 175 ft, or 
about the same as the depth of the exten- 
sively mined No. 5 coalbed. This agrees 
with theoretical models that predict low- 
frequency waves from strongly reflecting 
near-surface horizontal layers. 

The existing geologic structure between 
the main part of Blanford and the active 
pit is another possible or contributing 
cause of the low-frequency blast vibra- 
tions. This region has a coal cutout or 
zone where the coal and other sedimentary 
rock beds are missing, replaced by fill 
characterized as sandy, gravelly drift. 
Vibrations propagating through such mate- 
rial often have abnormally low frequency. 
Such a medium could also explain the 



rapid vibratiion amplitude attenuation 
observed between 407 and 807 ft (fig. 5). 
Blast designs are also significant. 
The widely adopted 8-ms charge-separation 
criterion appears not to apply for this 
low-frequency site, as it was also pre- 
viously found suspect in a 1979 study by 
Wiss of large-scale surface coal mine 
blasting (8). Vibration frequency char- 
acteristics appear to reflect periodici- 
ties in the blast design timing, partic- 
ularly close to the shot. The larger 
casting blasts produced the clearest 4-Hz 
surface waves. By contrast, the more 
complex echelon blasts produced the larg- 
est vibration amplitudes for a given 
charge weight per delay. Results suggest 
that longer delays be used between 
charges to prevent constructive addition 
of vibration amplitudes for such low- 
frequency cases. 



REFERENCES 



1. Siskind, D. E., M. S. Stagg, J. W. 
Kopp, and C. H. Dowding. Structure Re- 
sponse and Damage Produced by Ground Vi- 
bration From Surface Mine Blasting. Bu- 
Mines RI 8507, 1980, 74 pp. 

2. Duvall, W. I., C. F. Johnson, A. V. 
C. Meyer, and J. F. Devine. Vibrations 
From Instantaneous and Millisecond- 
Delayed Quarry Blasts. BuMines RI 6151, 
1963, 34 pp. 

3. Anderson, D. A. , S. R. Winzer, and 
A. P. Ritter. Blast Design for Optimiz- 
ing Fragmentation While Controlling Fre- 
quency of Ground Vibration. Paper in 
Proceedings of Eighth Annual Conference 
on Explosives and Blasting Technique, ed. 
by C. J. Konya (New Orleans, LA, Jan. 31- 
Feb. 4, 1982). Soc. Explos. Eng. , Mont- 
ville, OH, 1982, pp. 69-89. 

4. Anderson, D. A., A. P. Ritter, S. 
R. Winzer, and J. W. Reil. A Method for 
Site-Specific Prediction and Control of 
Ground Vibration From Blasting. Paper in 
Proceedings of First Mini-Symposium on 
Explosives and Blasting Research, ed. by 
C. J. Konya (San Diego, CA, Jan. 31-Feb. 
1, 1985). Soc. Explos. Eng., Montville, 
OH, 1985, pp. 28-43. 



5. Reil, J. W. , D. A. Anderson, A. P. 
Ritter, D. A. Clark, S. R. Winzer, and 
A. J. Petro. Geologic Factors Affecting 
Vibrations From Surface Mine Blasting 
(contract HO222009, Vibra-Tech Eng., 
Inc.). BuMines OFR 33-86, 1985, 204 pp.; 
NTIS PB 86-175858. 

6. Anderson, D. A., S. R. Winzer, and 
A. P. Ritter. Synthetic Delay Versus 
Frequency Plots for Predicting Ground Vi- 
bration From Blasting. Paper in Proceed- 
ings of the Third International Confer- 
ence on Compute r-Aided Seismic Analysis 
and Discrimination (Catholic University, 
Washington, DC, June 16-17, 1983). IEE 
Computer Society Press, Silver Spring, 
MD, 1983, pp. 70-74. 

7. Wiss, J. F. , and P. Linehan. Con- 
trol of Vibration and Blast Noise From 
Surface Coal Mining (contract JO255022, 
Wiss, Janney, Elstner and Associates, 
Inc.). Volume II. BuMines OFR 103(2)- 
79, 1978, 280 pp.; NTIS PB 299 888. 

8. Kopp, J. W. , and D. E. Siskind. 
Effects of Millisecond-Delay Intervals on 
Vibration and Airblast From Surface Coal 
Mine Blasting. BuMines RI 9026, 1986, 
44 pp. 



72 

9. Siskind, D. E. , V. J. Stachura, Geophy. Prospect., v. 5, 1957, pp. 371- 

and M. J. Nutting. Low-Frequency Vibra- 380. 

tions Produced by Surface Mine Blasting 11. Boscardin, M. D. Building Response 

Over Abandoned Underground Mines. Bu- to Excitation-Induced Ground Movements. 

Mines RI 9078, 1987. Ph.D. Thesis, Univ. IL, Urbana-Cham- 

10. O'Brien, P. N. S. Multiply-Re- paign, IL, 1980, 279 pp. 
fleeted Refractions in a Shallow Layer. 



73 



COMPUTER MODELING OF ROCK MOTION 
By Stephen A. Rholl 1 



ABSTRACT 



A computer model of rock motion due to 
blasting is presented. The code, CAROM, 
was developed at Sandia National Labora- 
tories (SNL) to predict rock motion and 
final rauckpile distribution. Researchers 
at the Bureau of Mines applied the code 



to simulate bench blasting during full- 
scale fragmentation tests at a nearby 
rock quarry. Results of the code are 
shown for the first 2 s following explo- 
sive detonation. 



INTRODUCTION 



A contributing factor to the overall 
efficiency of surface raining is the final 
muckpile distribution of the blasted rock 
fragments. This is especially true in 
mining operations utilizing overburden 
casting to reduce stripping costs. As 
the overburden-to-coal ratio increases, 
it is essential, in many operations, that 
as much fragmented rock as possible be 
thrown into the pit or onto the spoil 
bank. 

To optimize overburden casting, mine 
operators most often vary powder factor, 
explosive type, drill patterns, and/or 
delay timing. These variations are nor- 
mally tested by trial and error based on 
the experience of blasting personnel. 
Unfortunately, this can often be time 
consuming and expensive. An alternative 
approach to the optimization problem, 



which can quickly analyze technical 
problems, involves computer-simulated 
blasting. 

Much of the early work to develop sim- 
ple and effective computer models to sim- 
ulate blasting was done by SNL. In 1983, 
researchers developed BUMP (_1_), 2 the 
first code to use very simple interaction 
laws to reduce computational times. A 
stabilized and improved code, CAROM (2^), 
was later introduced. Both codes were 
written to study rock motion in oil 
shale crater tests. 

The Bureau of Mines initiated contract 
JO245011 with SNL to modify these codes 
for use in modeling bench blasting. Sub- 
sequently CAROM was modified and used in 
blast design research. SNL staff also 
provided appropriate values for input 
parameters to execute the code. 



MODEL DESCRIPTION 



The CAROM code is a two-dimensional 
distinct-element code. That is, a group 
of distinct elements is used to describe 
the system of rock fragments undergoing 
motion. The shape of these elements is 
in theory arbitrary, but circles are most 
often used to simplify time-of-collision 
calculations. The size of the elements 
is unrestricted, and often a single 
element is used to represent a group of 
fragments. Studies by Gorhara-Bergeron 
(2) have shown that the size of the 
material elements does not have a strong 

i 

'Research physicist, Twin Cities Re- 
search Center, Bureau of Mines, Minneap- 
olis, MN. 



influence on the calculated material mo- 
tion for modeling oil shale cratering 
experiments. 

The initial configuration of the ele- 
ments used in the CAROM code is specified 
by the user. In addition, the initial 
conditions (both the velocity and accel- 
eration of each element) must likewise be 
specified. CAROM simply lets the ele- 
ments move about until the code detects a 
collision between two elements. The col- 
lisions are handled by the theories of 
classical mechanics, namely, conservation 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



74 



of linear momentum and Newton's collision 
rule. The collisions themselves can be 
treated elastically or inelastically , al- 
though the latter is more often assumed. 

Because CAROM calculates the time the 
next collision will occur, the code is 
inherently stable. When a collision is 
detected, CAROM redistributes the linear 
momentum of the elements and then repeats 
the process. Collisions with a fixed 
boundary (or boundaries) are also permit- 
ted and treated in a similar way. 



Although it is possible to include 
friction in the calculations, CAROM usu- 
ally assumes that the coefficient of 
friction between the elements is zero. 
However, it should be noted that at the 
boundaries representing the pit floor, 
the elements are not allowed to slide. 
Too much sliding was deemed unrealistic, 
and therefore the code assumes infinite 
friction between the elements and the pit 
floor boundaries. 



MODEL APPLICATION 



The CAROM code was applied by Bureau 
researchers to simulate rock, motion dur- 
ing bench blasting at a local quarry. A 
two-dimensional discrete rock model de- 
scribing the 22-ft highwall is shown in 
figure 1. A series of 33 elements, each 
with a radius of 1 ft, represent the 
highwall. The explosive column is also 
indicated in the figure. 

After detonation of the explosives, 
shock waves are generated in the sur- 
rounding rock, which not only fracture 
the rock, but also impart momentum caus- 
ing unconfined rock to begin motion. In 
addition, large gaseous pressures are 
created within the boreholes that exert 
forces on the elements causing them to 
accelerate. 



10 



Z 5 



-i 1 1 1 1 1 1 1 1 1 1 1 1 r- 

NOTE. — Circles represent rock elements. 




I 



_i_ 



_i i i_ 



5 10 

HORIZONTAL DISTANCE, ra 



15 



FIGURE 1 .—Highwall representation by discrete rock elements. 
Heavy line on left indicates explosive column. 



Research at SNL has shown that a dynam- 
ic finite-element code, DYNA2D (3), can 
be used to describe the explosive detona- 
tion, rock fragmentation, borehole pres- 
sures, and initial rock motion. The in- 
formation calculated by DYNA2D can be 
used as the input necessary for the CAROM 
code. Both the momentum imparted to the 
rock fragments and the blast pressures 
due to the explosive gases are chosen 
based on DYNA2D calculations. DYNA2D 
cannot be used to model the entire blast 
because the code is a continuous repre- 
sentation, meaning it cannot describe 
material that breaks apart. 

Beside the initial conditions, further 
assumptions were necessary prior to run- 
ning the CAROM code. For the quarry 
tests, the size of each element was fixed 
and not allowed to change. This meant 
that elements were not permitted to break 
up into smaller fragments. Also, to sim- 
plify calculations, the elements were not 
allowed to roll and therefore, could not 
possess any angular momentum. The co- 
efficient of restitution for collisions 
between elements was chosen to be a con- 
stant for all of the elements. A value 
of 0.01 was assumed based on research at 
SNL. This implies that all of the col- 
lisions are treated as nearly perfectly 
inelastic. One final assumption deals 
with the collision criterion. Based on 
work at SNL, a collision is defined to 
occur when the distance between the cen- 
ters of two elements equals 1.9 ft. 

The quarry tests were conducted as sin- 
gle-row bench blasts to study the effect 



75 



of delay timing on fragmentation. Four of delay timing between holes in a single 
blastholes were fired in each of six 
full-scale shots. Because CAROM is at 
the moment only a two-dimensional code, 
it cannot be used to predict the effects 



row. However, once a three-dimensional 
code becomes available, these studies can 
be done. 



RESULTS 



All of the test shot 
with two high-speed came 
500 frames per second to 
al rock motion. Analys 
ing films yielded data p 
for confirmation of the 
primary interest was the 
velocity of the bench 
This was observed in a 
about 15 m/s. 



s were monitored 

ras operating at 

record the actu- 

is of the result- 

roviding a basis 

CAROM code. Of 

initial outward 

face material. 

11 tests to be 



The output from the CAROM code is shown 
in figures 2-11 for the first 2 s of rock 
motion at 200-ms intervals. The average 
velocity for any element can easily be 
calculated from the figures. The element 
representing the leading edge of the 
blasted rock face, as shown in figure 2, 
has an average velocity determined to be 
16.5 m/s, which is in excellent agreement 
with the cinematographic data. 




10 



5 10 

HORIZONTAL DISTANCE, m 

FIGURE 2.— Predicted rock motion at time equals 0.2 s. 



15 



p; s 







-J 1 1- 



NOTE. — Circles represent rock elements. 




o£o 



qcQ d : 



QVP 



o. 



o 



Cr 



_Q 



O 



o 



5 10 

HORIZONTAL DISTANCE, m 

FIGURE 3.— Predicted rock motion at time equals 0.4 s. 



10 



-i 1 r- 



: 



NOTE. — Circles represent rock elements. 




o° 
oo 

or£b . 9) . on? . o ■ ° o. 



10 



5 10 

HORIZONTAL DISTANCE, m 

FIGURE 4.— Predicted rock motion at time equals 0.6 s. 



15 





1 1 | IT" — 1 1 

NOTE. — Circles represent 


i i i » — i — 

rock elements. 


D 








- 






0°C 


o 


o 


c? ' 






oooo 


tt- 


8tfb 


Oo . . 




5 10 

HORIZONTAL DISTANCE, m 

FIGURE 5.— Predicted rock motion at time equals 0.8 s. 



15 



76 



10 



"~i 1 T r~ 



~i 1 I 1 1 1 1 1 1 1 1 1 ~| r~ 

NOTE. — Circles represent rock elements. 




n i no oorro 



o o 



5 10 15 

HORIZONTAL DISTANCE, m 

FIGURE 6.— Predicted rock motion at time equals 1.0 s. 



20 









1 i 


NOTE. — 


| i i i i | i i i i 
Circles represent rock elements. 


E 

UJ* 
U 

< 

in 
a 


5 


- 






- 


< 

u 

01 

u 
> 




Eb° 


O \6 


A 


o 

o °: 

i on , , i , , , , 



5 10 15 

HORIZONTAL DISTANCE, m 

FIGURE 7.— Predicted rock motion at time equals 1.2 s. 



20 



10 



p; 5 



. . , , , 


i 


i i i 


1 ' ' 


"i 1 i i i i i 1 i i i i 

NOTE. — Circles represent rock elements. 


: 




°or 


-\ 


i i i i 
O 


3r9r>. o . 


rxx 


^ 


), on 


o , ° - 

i i i i i i i i i i j i 



5 10 15 

HORIZONTAL DISTANCE, m 

FIGURE 8.— Predicted rock motion at time equals 1.4 s. 



20 



25 



77 



10 



E 5 



1 1 1 


I 1 1 1 


i i i i , 


1 1 | 1 1 1 I ■ 1 ■ T ' "1 1 1 

NOTE. — Circles represent rock elements. 


2*0 




Cl 


o 


3¥&. 


ooSxr 


rm? oo 


i i v si i i i i 1 i i i KJ 



5 10 15 

HORIZONTAL DISTANCE, m 

FIGURE 9.— Predicted rock motion at time equals 1.6 s. 



25 



10 



-i 1 1 r 



~i 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 r- 

NOTE. — Circles represent rock elements. 



5 5 




o_ 



_Q 



cxPr&W) 



i o oo , , o 



o 



_l 1 L 



n. 



5 10 15 

HORIZONTAL DISTANCE, m 

FIGURE 10.— Predicted rock motion at time equals 1.8 s. 



25 



10 



-i 1 1 r~ 



-t 1 1 r 



1 ' ' ' <~ , 

NOTE. — Circles represent rock elements. 



S 5 




nrxxr9^ff 



ooo , ,o 



_i i i i_ 



o, o 



5 10 15 

HORIZONTAL DISTANCE, m 

FIGURE 1 1 .—Predicted rock motion at time equals 2.0 s. 



20 



25 



78 



In the fragmentation test shot utiliz- 
ing 4.0 ms/ft of burden as the delay tim- 
ing, the maximum rock throw was measured 
to be 30 m. The CAROM code predicted, as 
shown in figure 9, at least one element 
thrown to about 24 m after 1.6 s of time 
passage in the calculations. CAROM is 
also capable of predicting vertical mo- 
tion of the bench top. However, in the 
series of results presented here no up- 
ward motion is observed. It is possible 
that CAROM still did indicate such mo- 
tion, but that it occurred prior to 
200 ms (fig. 2). At this time CAROM 
shows the surface material falling down- 
ward under the influence of gravity. The 
high-speed films confirmed that there was 



little or no vertical rock motion in the 
field tests. 

One of the deficiencies of the code ap- 
pears to be its inability to predict the 
profile of the rock motion of the bench 
face. As shown in figure 2, CAROM pre- 
dicts almost the entire face to break off 
as a slab. It is likely that this is due 
to improper balancing of the input param- 
eters. That is, there is too much empha- 
sis on the imparted momentum due to the 
shock wave impulse, and not enough empha- 
sis on the accelerations due to the 
forces created by the pressures of the 
explosives gases. Research continues to 
develop a step function approximation for 
the high-pressure gases. 



CONCLUSIONS 



The computer code CAROM, which models 
rock throw in blasting, has been briefly 
discussed. The code was used to simulate 
rock motion at a limestone quarry, and 
the results of the analysis have been 
presented. The code accurately predicted 
the velocity of burden movement and the 



maximum throw of the rock fragments. 
CAROM is still being developed, and many 
improvements are being implemented. The 
objective continues to be to provide 
mine operators with accurate predic- 
tions of rock motion and final muckpile 
distribution. 



REFERENCES 



1. Scharaaun, J. T. An Engineering 
Model for Predicting Rubble Motion During 
Blasting. Paper in Proceedings of the 
9th Conference on Explosives and Blasting 
Technique, ed. by C. J. Konya. (Dallas, 
TX, Jan. 31-Feb. 4, 1983). Soc. Explos. 
Eng. , Montville, OH, 1983, pp. 199-222. 

2. Gorhara-Bergeron, E. Rock Motion 
Modeling of Oil Shale Cratering Experi- 
ments. Paper in Research and Engineering 



Applications in Rock Masses. Proc. 26th 
U.S. Symp. on Rock Mechanics, Rapid City, 
SD, June 1985, Soc. Min. Eng., 1985, 
pp. 151-159. 

3. Hallquist, J. 0. User's Manual for 
DYNA2D - An Explicit Two-Dimensional Hy- 
drodynaraic Finite Element Code With 
Interactive Rezoning. Lawrence Livermore 
Nat. Lab., Livermore, CA, UCRL-52997, 
1982, 109 pp. 



79 



INFLUENCE OF BLAST DELAY TIME ON ROCK FRAGMENTATION: ONE -TENTH-SCALE TESTS 
By Mark S. Stagg 1 and Michael J. Nutting 2 



ABSTRACT 



The Bureau of Mines is studying blast 
delay timing influences on rock frag- 
mentation in a series of tests that 
started in 3-ft concrete blocks and in- 
cludes reduced-scale and full-scale bench 
blasts. This paper reports on the 
reduced-scale tests. In a 45-in-high 
dolomite bench, 18 single-row blasts were 
fired with 15-in burdens. Spacings were 
21 and 30 in. Delay intervals ranged 
from to 45 ms, equivalent to to 36 
ms/ft of burden. Each shot was instru- 
mented for strain and pressure for both 
in situ dynamics and interactions between 
blastholes. All fragmented rock was 
screened. 

The finest fragmentation occurred at 
blasthole delay intervals of 1 to 17 



ms/ft of burden. In this range, stress- 
wave-induced strains interacted with 
longer lasting gas-pressure strains from 
earlier holes. Coarse fragmentation re- 
sulted from short delays (<1 ms/ft), 
where breakage approached presplit con- 
ditions with a major fracture between 
blastholes and large blocks in the burden 
region. Coarse fragmentation also re- 
sulted from long delays (>17 ms/ft), with 
explosive charges acting independently. 
The broad acceptable range provides blast 
design tools for a variety of purposes, 
including optimum muckpile displacement 
and vibration control. 



INTRODUCTION 



The explosives industry is developing 
and testing delay blasting caps of im- 
proved accuracy. Precise delays have 
been cited as factors in controlling 
blast vibration amplitude and frequency 
and improving fragmentation (1_~_2)» 3 How- 
ever, data on complete fragmentation as- 
sessment of shots initiated with precise 
days is limited. As part of its blasting 
research program, the Bureau of Mines is 
examining the influence of timing inter- 
vals by completely screening the blast- 
induced rock. Three- and four-hole shots 
have been conducted in concrete blocks in 
the laboratory and at reduced (45-in 
bench) and full (22-ft bench) scale in 
the field. Tests thus far have mostly 
been concerned with the effect of de- 
lay time on fragmentation and on the 

Civil engineer. 
2 Geophysicist (now with Philip R. Ber- 
ger & Associates, Inc., Warrendale, PA). 

Twin Cities Research Center, Bureau of 
Mines, Minneapolis, MN. 

■^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



interaction between shotholes. Initial 
testing in the laboratory provided an 
effective means for establishing a meth- 
odology of controlled experimentation. 
The tests at reduced scale in the field 
provided experience in fragmentation 
assessment techniques and results that 
could be used to optimize the expensive 
full-scale field tests. The full-scale 
field tests are currently in progress. 
This paper discusses the reduced-scale 
field tests and results. 

The reduced-scale field tests were con- 
ducted at the University of Missouri's 
Experimental Mine in Rolla. This site 
was chosen because of its accessibility 
and geology, and the cooperation avail- 
able from the University. Furthermore, 
the results of previous research con- 
ducted at the mine on blast design and 
fragmentation were reported in several 
theses (3-_5)» These studies investigated 
various design factors affecting fragmen- 
tation, such as coupling, initiation se- 
quence, primer location, and airgap, anc 1 
provided a comparison with the Bureau 
test results. 



80 



SITE 



The experiment was conducted in a 45-in 
bench of dolomite, part of the Jefferson 
City Formation. The rock is of irregular 
grain size, 10 pet calcite, and thick 
bedded, with a specific gravity of 2.65 
and corapressional and shear velocities of 
14,800 and 8,100 ft/s, respectively (_5). 
Bureau researchers verified these values 
by in situ seismic measurements behind 
the blastholes, finding 14,700 ft/s (com- 
pressional) and 8,100 ft/s (shear). 

The reduced-scale tests were designed 
to be geometrically proportional to 



typical full-scale bench blasts with di- 
mensions about 10 pet of full scale. 
However, at reduced scale, rock struc- 
tures such as bedding and jointing are 
exaggerated and can have an unrealistic 
effect. The massive dolomite at the Rol- 
la site provided a good medium for the 
reduced-scale testing since only three 
bedding planes were in the 45-in bench 
and no jointing was observed within any 
of the test shots. 



PROCEDURES 



In order to prepare the pit and bench 
for the tests, development work was nec- 
essary to provide consistent geometry of 
the working faces. Horizontal holes were 
drilled 45 in above the pit floor to give 
the proper bench height, and vertical 
holes were drilled to obtain a vertical 
face. For each test shot, an open end 
at an angle of approximately 135° was 
formed, as shown in figure 1. To form 
the single-row pattern, three shotholes 
were drilled to 50-in depth, including 
5 in of subdrill. The burden was held 
constant at 15 in, and spacings were 21 
and 30 in, spacing-to-burden ratios (S/B) 
of 1.4 and 2.0, respectively. All faces 
were cleaned with a scaling bar and blown 



Hole I Hole2 Hole3 
4« 21 or 30"-—+ 21 



-135° N-. AT 



--o-- 



or 30" — »| 



'"* "' \lk. Lit Ul— 



-<K 



LiL .., ,s, », /// (// 



Wi — 75T-^ 5 




FIGURE 1.— Shot pattern and blast design for reduced-scale 
tests. 



with air to remove loose material, and 
the entire area was swept clean prior to 
each shot. 

Each hole contained 144 g of 60 pet 
extra-high-density dynamite tamped into a 
1/2-in-ID plastic tube 40 in long. Each 
charge was bottom-primed and initiated by 
an exploding bridgewire detonator (EBW). 
Complete coupling was assured by placing 
the charge into grout-filled holes. The 
initiation system for the EBW's consisted 
of a power supply, firing module, and 
digital delay generator with a firing 
accuracy of 0.0025 pet times the delay 
time or ±50 ns, whichever is greater. 

Most shots were instrumented with 
dynamic-strain and pressure gauges (figs. 
2-3). The strain gauges were a six-com- 
ponent type, built after a design by Reed 
(6^) as modified by Anderson (7). These 
gauges were grouted into the burden re- 
gion at various locations between the 
boreholes. Pressure gauges were placed 
either above the strain gauges or in in- 
clined holes in the face, which were 
filled with a water-revert mixture. The 
pressure gauges were of two types, carbon 
resistors dipped in liquid tape (insulat- 
ing coating) and Navy-built tourmaline 
gauges in an oil-filled boot. Also, a 
fiber optic system was used to measure 
detonation velocities. Data were re- 
corded on a 28-channel Wide Band I (80 
kHz) recorder and a digital oscilloscope, 
with a 0.5-ys response rate. 

To contain flyrock and minimize second- 
ary breakage, the entire shot was covered 



81 





IS 






4 



-5 



FIGURE 3. — Carbon resistor (left and right) and tourmaline 
(center) pressure gauges. 



FIGURE 2.— Stages of assembly of six-component strain 
gauge. 

with a blasting mat held in place with 
timbers and anchoring cables (fig. 4). 
The area in front of the shots was cov- 
ered with a plastic sheet to aid in .iden- 
tifying blasted material. Flyrock went 
beyond the sheeting for only a few shots. 
This rock was identified when possible 
and included with the muckpile. 

Screening of the muckpile began immedi- 
ately after each shot. All fragments 
3 in and larger were sized and weighed in 
the pit. The pit size fractions were 
minus 3, plus 3, minus 6, plus 6, minus 
12, and plus 12 in. Material passing the 
3-in screen was loaded into containers, 
removed from the pit, and mechanically 
shaken through screens. These size frac- 
tions were plus 1-1/2 minus 3, plus 3/4 
minus 1-1/2, plus 3/8 minus 3/4, plus 
3/16 minus 3/8, and minus 3/16 in. From 
the weight of each fraction, its percent- 
age of the total muckpile was calculated. 




FIGURE 4.— Pit area with blasting mat and timbers covering 
test shot. 



ANALYSIS AND RESULTS 



A total of 24 blasts were detonated. 
Two were development shots. Of the re- 
mainder, nine at 21-in spacing and 
nine at 30-in spacing were completely 
screened, three misfired, and one was a 



single-hole test shot. Delay intervals 
ranged from simultaneous to 36.0 ms/ft of 
burden. Table 1 lists the reduced-scale 
(RS) shots. 



82 



Since these tests were designed to de- 
termine the effect of delay time on frag- 
mentation, i.e., the interaction between 
shotholes, it was decided to identify any 
overbreak from each shot and to size and 
weigh it separately from the rauckpile. A 
third of the shots produced no back- or 
end-overbreak. There was no obvious cor- 
relation between spacing or timing and 
those shots producing overbreak, which 
was generally oversized and averaged 10 
pet of the total blasted rock weight. 
Because the overbreak skewed the particle 
size distributions at the high end and 



came from outside the shot pattern, it 
was removed from the corresponding size 
fraction and not included in the analy- 
sis of fragmentation. Table 2 lists the 
sieve weights for the shots. 

The cumulative percent passing versus 
sieve size is plotted in figure 5 for 
four shots at each spacing, covering the 
range of delay times tested. The mate- 
rial that passed through the 24-in sieve 
would in most cases have passed through a 
much smaller sieve, even down to 13 in. 
Since the largest size piece was not mea- 
sured and 24 in is too large, the 24-in 



TABLE 1. - Reduced-scale (RS) test shots 





Spacing, 
in 


Delay 


time 


Shot 1 


Spacing, 
in 


Delay 


time 


Shot 1 


Between 


ms/ft of 


Between 


ms/ft of 






holes, ms 


burden 






holes, ms 


burden 




30 


20.0 


16.0 




30 


1.25 


1.0 




30 


7.5 


6.0 




30 








RS-5 


30 


5.0 


4.0 




30 


45.0 


36.0 


RS-6 


30 


20.0 


16.0 




30 


2.5 


2.0 


RS-9 


21 










21 


21.0 


16.8 




21 


30.0 


24.0 




21 


1.75 


1.4 


RS-11 


21 


14.0 


11.2 




21 


.4375 


.35 




21 


3.5 


2.8 




21 


8.75 


7.0 


RS-13 


30 


30.0 


24.0 




21 


45.0 


36.0 



] Not listed: Shots RS-1 and RS-2 (development shots), RS-7, RS-8, and RS-20 (mis- 
fires), and RS-23 (single-hole shot). 



TABLE 2. - Weight of rock fragments at various screen sizes, pounds 







+ 3/16 


+ 3/8 


+ 3/4 


+1-1/2 


+ 3 


+ 6 


+ 12 


Screen size, in... 


-3/16 


-3/8 


-3/4 


-1-1/2 


-3 


-6 


-12 


-24 


RS-3 


208 
172 
215 


243 
243 
237 


268 
263 
258 


535 
570 
498 


639 
675 
523 


1,389 
1,987 
1,628 


1,607 
2,022 
1,941 


241 


RS-4 


241 


RS-5 


159 


RS-6 


188 


216 


257 


466 


596 


1,656 


1,745 


117 


RS-9 


196 
232 
305 


182 
259 
318 


196 
320 
391 


356 
552 
587 


421 
631 
572 


1,197 
1,372 
1,149 


1,173 

2,100 

800 


1,415 


RS-10 


172 


RS-11 







233 


255 


298 


557 


639 


1,521 


1,147 





RS-13 


238 


238 


276 


452 


516 


1,550 


1,857 


520 


RS-14 


217 


240 


253 


541 


612 


1,911 


1,718 


266 


RS-15 


167 
244 


182 
250 


208 
246 


466 
471 


521 
512 


1,657 
1,607 


2,092 
2,019 


462 


RS-16 


536 




242 


247 


309 


522 


630 


1,801 


1,546 


446 


RS-18 


246 
202 


275 
234 


342 
302 


522 
548 


590 
635 


1,197 
1,219 


876 
911 










RS-21 


203 
268 


194 
235 


198 
258 


377 
454 


449 
501 


1,132 
1,131 


1,507 
1,022 


688 




143 




71 


60 


61 


123 


138 


364 


550 


178 




295 


249 


270 


486 


541 


1,060 


1,110 


422 



83 



-i — i — i — n 



KEY 
o RS-9, ms/fl 
o RS-19, 1.4 ms/tt 
» RS-18,16.8 ms/ft 
» RS-24, 36.0 ms/fl 
-Extrapolation of data, see text 




i — i — i i i i 



30-in spacing 



KEY 
o RS-15, ms/ft 
o RS-14, 1.0 ms/tt 
« RS-6, 1 6.0 ms/ft 
v RS-24, 36.0 ms/ft 




1.0 
SIEVE SIZE, in 

FIGURE 5.— Particle size distribution for shots at 21- and 30-in 
spacings. 



point has been omitted from the curves, 
except that it was used to obtain the 80- 
pct-passing value for shot RS-9, as shown 
in figure 5. The curves in figure 5 are 
representative of the results, which 
showed that a dramatic improvement, a 20- 
to 50-pct reduction in average size, oc- 
curred at delays of ~1.0 to 17.0 ms/ft 
compared with simultaneous initiation. 
At delays longer than 17.0 ms/ft, the 
average size increased ~20 to 50 pet. 

The 20-, 50-, and 80-pct passing sizes 
versus delay period, shown in figure 6, 
were determined for all the tests from 
percent-passing curves similar to those 
shown in figure 5. The delay period had 
little effect on the size of fragments 
in the 20-pct-passing fraction, except 
that they were slightly coarser at the 
simultaneous shots. At both spacings, 
the 50- and 80-pct-passing fractions show 




DELAY, ms/ft of burden 

FIGURE 6.— Size at 20, 50, and 80 pet passing versus delay 
period for shots at 21- and 30-in spacings. 



that delays of 1 to 17 ms/ft produced 
the finest fragmentation. Poorest or 
coarsest fragmentation was observed for 
simultaneous shots and for shots with 
delay intervals >24 ms/ft, although the 
differences at the longer delays were 
smaller with the 30-in spacings. The im- 
proved fragmentation obtained for most of 
the shots at 21-in spacings, as compared 
with 30-in spacings, was expected and was 
due in part to a higher powder factor. 
However, the coarsest fragmentation re- 
sulted at the very short delay times for 
the 21-in spacing tests, even though 
there was a higher powder factor. 

To quantify the test results and ulti- 
mately develop an equation to predict 
fragmentation, it is necessary to develop 
a mathematical description of the cumu- 
lative percent-passing data. Previous 
researchers have used regression analy- 
sis to fit observed blast fragmentation 
data to logarithmic, power (8) , Gaussian 
(2), and Weibull distributions (9-10). 
Analysis-of-variance tests can then be 
used to determine the statistical signif- 
icance of the effect of delay time on the 
cumulative size distribution functions. 
These tests determine whether a single 
regression line can be used to represent 
the combined results of several shots. 
If pooling or combining the results tests 
positive, then there exists a statistical 
inference that the delay time has not 
significantly affected fragmentation. 



84 



Statistical tests were run on several 
regression line fits to the data, and the 
correlation coefficients (R) for the var- 
ious distributions are given in table 3. 
The data shown in figure 6 suggest that 
the finer size material was less affected 
by delay time. An examination of the 
weights of the material passing the 3-in 
sieve found minimal variability, as shown 
in figure 7, except for shots at delays 
less than 1.0 ms/ft. Excluding these 
shots and an extremely high (2,170 lb) 
outlier value, the weight of material 
passing the 3-in sieve ranged from 1,716 
to 1,994 lb and 1,720 to 1,950 for the 
21- and 30-in spacing shots, respective- 
ly. The nearly constant weight of finer 
material suggests that a fractured zone 
extends around the borehole, which, as- 
suming a cylindrical nature, would have 
a radius of 10 in. This equates to 40 
explosive radii, within the range of 
the damage zone predicted by Siskind (11- 
12) , Olson (13), and others in terms of 
charge radii (i.e., 20 to 40 radii). As 
shown in figure 8, the fines best fit a 



log-normal distribution. Analysis-of- 
variance tests were run for all finer 
size data from delays of 1 to 36 ms/ft, 
and one curve could be used to represent 
the data at both spacings. For shots at 
delays of less than 1 ms/ft, the weight 
of fines was reduced. 

The correlation coefficients in table 3 
show that a Gaussian distribution for 
the coarser size data usually produced 
the best fit. The Gaussian, power, and 
Weibull fits to the coarse size data are 
shown in figure 9 for shots RS-3 and 
RS-13. It was observed that the 1-1/2-in 
sieve material also fit the Gaussian dis- 
tribution, and this material was included 
in the regression analysis of table 3. 
The coarse material is assumed to be gen- 
erated primarily between the boreholes 
outside the extended fractured zone that 
exists around the borehole. 

An attempt to pool all of the data us- 
ing a Gaussian distribution indicated (at 
a 95-pct-conf idence level) that the delay 
time was indeed significant. Further 
analysis-of-variance tests were conducted 



TABLE 3. - Correlation coefficients for Weibull, 
power, and Gaussian distributions 



Shot 


Weibull 


Power 


Gaussian 




A 1 


B 2 


A 1 


B 2 


B 2 


RS-3 


0.9923 
.9906 
.9833 
.9856 
.9961 
.9859 
.9966 
.9958 
.9893 
.9893 
.9895 
.9872 
.9832 
.9966 
.9978 
.9921 
.9904 
.9956 


0.9861 
.9858 
.9758 
.9815 
.9936 
.9720 
.9852 
.9845 
.9847 
.9868 
.9848 
.9830 
.9906 
.9875 
.9909 
.9887 
.9862 
.9921 


0.9963 
.9961 
.9972 
.9874 
.9978 
.9966 
.9876 
.9948 
.9973 
.9972 
.9982 
.9971 
.9970 
.9902 
.9913 
.9981 
.9966 
.9968 


0.9980 
.9944 
.9941 
.9953 
.9942 
.9980 
.9956 
.9941 
.9946 
.9923 
.9948 
.9940 
.9928 
.9933 
.9932 
.9970 
.9957 
.9980 


0.9999 


RS-4 


.9995 


RS-5 


.9993 


RS-6 


.9998 


RS-9 

RS-10 


.9837 
.9979 


RS-11 


.9996 


RS-12 


.9992 


RS-13 


.9988 


RS-15 


.9981 
.9993 


RS-16 


.9992 


RS-17 


.9945 


RS-18 


.9998 


RS-19 


1.0000 


RS-21 


.9980 


RS-22 


.9994 




.9964 



'All data 
the +12-, 

2 0nly dat 
100-pct po 



points were u 
24-in data poi 
a at 1-1/2-, 3 
int was also e 



sed in regression analysis except 

nt. 

-, 6-, and 12-in sizes were used; 

xcluded. 



85 



6400 



4.800 m 



3.200 



1,600 




RS-9 RS-21 RS-19 RS-12 RS-22 RS-11 RS-1B RS-10 RS-24 

D. 35 1.4 2.8 7.0 11.2 16.8 24.0 36.0 



6.400 



4000 



3.200 



1.600 




RS-15 RS-14 RS-17 RS-5 RS-4 RS-3 RS-6 RS-13 RS-16 

1.0 2.0 4.0 6.0 16.0 16.0 24.0 36.0 

SHOT NUMBER AND DELAY TIME, ms/ft 




KEY 

Fraction 
,3 size, in 

■•■12 



*6 
-12 





'Mm^taa 




►3/2 
-3 



•■3/4 
-3/2 



► 3/8 
-3/4 



+3/ 16 
-3/8 



"6 

-12 



►3/2 
-3 



►3/4 
-3/2 



► 3/8 
-3/4 



► 3/16 
-3/8 



-3/16 



FIGURE 7.— Weight retained on each size fraction for shots at 21- and 30-in spacings. 



86 




80 
70 
60 
50 
40 
30 
20 



1-36 ms/ft 

21- and 30- in spacing 




'32 



FIGURE 8.— Log-normal distribution for material passing 1 Vi-in 
sieve size. 



to see if certain shots could be pooled 
to form one regression line. For exam- 
ple, the 21-in shots at 11.2 and 16.8 
ms/ft could be combined (i.e., there was 
no effect due to delay time). The 30-in 
shots at 1,2, 4, and 6 ms/ft could also 
be combined. Figure 10 is a plot of the 
50- and 80-pct-passing value determined 
from the Gaussian distributions. A hori- 
zontal line is drawn for delays that 
could be pooled into one regression line. 
Although not shown in the figure, the 
single-hole fragmentation distribution 
pooled with the 30-in spacing curves 
at delays of 24 and 36 ms/ft. Apparent- 
ly, firing holes at delays of >24 ms/ft 
can be considered as firing single-hole 
shots. It is noteworthy that the dis- 
tance from the corner to the first hole 
was 21 in (fig. 1), but the fragmentation 
data pooled with the 30-in spacing tests. 
The single-hole test reflected the 30-in 
results because at 21 in, the first 
hole's breakout angle, >135°, reduced the 
burden distance for subsequent holes, 



97 
95 
90 

80 
70 
60 
50 
40 
30 
20 



1 1 r 

a RS-3, 16.0 ms/ft 
O RS-13, 24.0 ms/ft 

Gaussian fit 




FIGURE 9. — Comparison of Gaussian, power, and Wiebull 
distributions. 



which improved fragmentation over that 
resulting from single-hole shot. 

The 50- and 80-pct curves of figure 6 
and 10 are quite similar except for the 
simultaneous shot at 21-in spacing, RS-9. 
The Gaussian distribution was not the 
best fit for this shot, and the regres- 
sion line predicts a higher value than 
the data suggest. 

As mentioned earlier, the University of 
Missouri's Experimental Mine has been 
used by several researchers (3-_5) to con- 
duct investigations of blast design and 
fragmentation. Where possible, these 
data have been compared with Bureau re- 
sults, as shown in figure 11. Since the 



87 



"I 1 F — I I I 



-| 1 

KEY 



1 ' ' ' I 




21-in spocinq 

50 pet passing 

a 80 pet passing 
30-in spacing 

• 50 pet passing 

■ 80 pet passing 

rapolotion of doto, see text 




-i — L 



DELAY, ms/fl of burden 



FIGURE 1 0.— Size at 50 and 80 pet passing versus delay period 
from Gaussian distribution fits to the 21- and 30-in spacing data. 




i io 

DELAY, ms/fl of burden 



FIGURE 11.— Comparison of reduced-scale data with results 
obtained by other researchers. 



shot designs for these tests were similar 
to those used by the Bureau, the fragmen- 
tation results compare very favorably. 
Other research, such as Bergmann's multi- 
hole tests in granite blocks (14) , showed 
a similar significant improvement in 
fragmentation as delay times increased 
from simultaneous to 1 ms/ft of burden. 
Winzer's tests in limestone blocks and in 
a small bench (15) resulted in a rela- 
tionship between delay time and fragmen- 
tation that correlates well with Bureau 
results. The character of the data in 
figure 11 is similar, showing substantial 



improvement in fragmentation as delays 
increase to 1 ms/ft, slightly coarser 
fragmentation between 6 and 7 ms/ft, and 
continued improvement to 10 ms/ft. 

Strain and pressure records obtained 
for the reduced-scale tests tend to con- 
firm the fracture development mechanisms 
observed and reported by Holloway in 
work done under contract to the Bureau 
(S0245046, 1986). Initially, a fracture 
zone develops around the borehole because 
of the development of radial fractures 
and fracturing caused by reflected stress 
waves. The radial fractures propagate at 
speeds down to 12 pet of the P-wave ve- 
locity (16). The fractured region for 
these tests appeared to coincide with the 
finer material zone, which was generated 
within about 40 charge radii. From pres- 
sure gauges installed in this zone, the 
velocity of explosive gases penetrating 
fractures was found to be approximately 
1,800 to 2,700 ft/s, as shown in figure 

12. The large impulsive signals on the 
records were due to the pickup of elec- 
trical noise from the EBW initiation sys- 
tem and were used to confirm the delay 
intervals. 

The P-wave velocities were determined 
from the arrival times and the distances 
from shotholes to gauges. The distance 
and arrival time measurements used to 
calculate the velocity of gas movement 
through the rock were adjusted to correct 
for the travel time of the bottom initi- 
ated explosive detonation (8,000 ft/s) to 
the height of the gauge. 

The borehole pressure and radial crack 
pressurization produce stresses in the 
material beyond the near fracture zone, 
and this leads to additional failure and 
extension of the radial cracks to the 
free face. The velocity of gas penetrat- 
ing fractures was estimated to be 146 
ft/s for shot RS-14, as shown in figure 

13. This velocity was determined by sub- 
tracting from the arrival time the travel 
time of the explosive to the gauge height 
and the time of gas penetration (1,800 
ft/s) out to 10 in. The remaining dis- 
tance to the gauge and the remaining time 
were used to determine the velocity. 

Strain data from the shots were pro- 
cessed into resultant principal strains, 
octahedral shear strain and dilatation 



88 



400 



300 



o 



m 

5j 200 
en 

CO 
UJ 

cr 
o_ 



Firing 
pulse, 
hole 3 



Firing 

pulse, Firing 
hole I pulse, 
hole 2 



t-Gas pressure, 1,760 ft/s 




(3) (§> 



//B£i[&^ \ltSA\&. 



7'A 



2-1 



r 



Gauge hole drilled 20°— j-o — r 
down from horizontal |fc' 

■ i 



h&s 




■(*=**■ 



FIGURE 12.— Pressure measured for shot RS-19. 



(_7_). Two of the principal strains are 
plotted in figures 14 through 17 for test 
shots with simultaneous and 1.4-, 2.0-, 
and 16.0-ms/ft delays. Shown on the rec- 
ords are the calculated P-wave and gas 
velocities determined from arrival times 
and shothole-to-gauge distances as dis- 
cussed earlier. The strain pulses in- 
creased in amplitude with decreasing dis- 
tances from shothole to gauge. These 
pulses correspond to the arrival of 
stress waves, which often arrived in- 
creasingly later in time as holes 2 and 3 
were fired. The observed decrease in 
corapressional (P) and shear (S) veloc- 
ities is probably due to flaws in the 
rock, such as fractures and cracks pro- 
duced by stress waves from an earlier 
shothole. 

The gas velocities observed on the 
strain records agreed with pressure gauge 
gas velocity observations, except for 
those in shot RS-2, where it is believed 



a major crack developed in line with 
holes 2 and 3, causing the observed gas 
velocity of ~2,000 ft/s. Excessive end- 
break was noted for this development 
shot. 

Optimum fragmentation occurred when a 
hole fired such that its stress wave in- 
teracted with the stress induced by the 
expanding gas pressurization from the 
previous hole. Shots RS-2 and RS-6 show 
a long-term strain believed to be induced 
by the late-arriving gas pressure inter- 
acting with the stress wave from hole 3. 
Similar measurements and observations 
have been made at full scale (2)» The 
interactions of strains induced by the 
stress waves and strains induced by gas 
pressure were not always observed for all 
shots of optimum delay (1 to 17 ms/ft), 
because the gauge is stressed only if 
near a pressurized crack and well-coupled 
to the rock. These interactions were 
also not observed for shots with delay 



89 



80t 



60 



o 

.a 



UJ 

a: 

co 40 

CO 
UJ 

rr 
a. 



20 



Gas pressure, 146 ft/s 
P-wave, 6,880 ft/s 



■ An * ~^ * v 



Firing 
pulse, 
hole I 



^,r^\ 




TIME, ms 



© © 



© 



I 



///,^/unZtn. 






!I5' 



1 

Gauge hole drilled */^~^~» 
20° down from -< 20 
horizontal _J L_ 




lli>//H> 



Firing Firing 
pulse, pulse, 
hole 2 hole 3 



FIGURE 13.— Pressure measured for shot RS-14. 



times outside the optimum range. Even 
though shot RS-19, shown in figure 15, 
does not show an interaction at the gauge 
location, pressure effects are still ob- 
served later in the record. Gas effects 
are not as apparent for the simultaneous 
shot, shown in figure 14. 

Fragmentation results at short delays 
(less than 1 ms/ft of burden) suggest 
that fracturing in the zone around the 
borehole must be completed before the 
next hole fires. Using 1,800 ft/s as the 



velocity for crack development in the 10- 
in zone around the borehole, plus the 
explosive detonation time (0.42 ms), the 
next hole should not be fired until after 
0.9 ms or 0.7 ms/ft. Enhanced cracking 
appears to last as long as the gas is re- 
tained. Gas velocities through the rock 
suggest the process will last up to 20 ms 
or 16 ms/ft, based on a 45° breakout an- 
gle and gas penetration velocities of 
1,800 ft/s for the first 10 in and 50 
ft/s for the next 11 in. 



90 



2 000 + /-Firing pulse, all holes 



,000 



c 
< 

DC 

w 



■* 



-1,000 



Shot material 



-2,000 




i&t&tetus 



FIGURE 14.— Principal strains measured for shot RS-9. Simultaneous shot. 



91 



10 '£ 



10 U," 1 



1,600 



1,200 



800 



400 



c 
< 



-400 



-800 



-1,200 



- 1,600 l 



— Firing pulse, hole I 



(D---Q-© .q 

Shot material 




I— P-wave, 13,100 ft/s 

I — Firing pulse, hole 2 

P-wave, 11,800 ft/s 

P-wave, 7,060 ft/s 

— Gas pressure, 60 ft/s 



Strain gauge, 25 in deep 




V/||h'~vA^w-*-*»/v* V \A/V»-**' v '' v ^^ 



6.0 



'—Firing pulse, hole 3 



FIGURE 15.— Principal strains measured for shot RS-19. Delay time, 1.4 ms/ft. 



92 



2,000 t 



1,000 



c 

— 
c 

Z 

< 
tr 
y- 
co 



-1,000 



-2,000 







/ \ ^ 

^-Strain gauge, 
25 in deep 



FIGURE 16.— Principal strains measured for shot RS-2. Delay time, 2.0 ms/ft. 



93 



2,800 



1,400 



c 

5. 



< 
a. 

(0 



-1,400 



P-wave 12,400 ft/s 



jj WAifr*»/yUW/^^ 



I — P-wave, 10,000 ft/s 



Gas pressure, 115 ft/s 




'rAfAiJffh 



P-wave, 7, 170 ft/s 



IV 



y' Y ^YV J <V''"^^~ r v>'>'r^"'vy'V'^ l n^ 



12 



Fire pulse, hole 2 
Fire pulse, hole I 




TIME, ms 



Fire pulse, hole 3 



Mk 



I 15 

N *■ 



II i e II 






-2,800 L 



®-x © <D — -y* 



Strain gauge, 25 in deep 
FIGURE 17.— Principal strains measured for shot RS-6. Delay time, 16.0 ms/ft. 




Shot material 




sgjyftMfeJ 



94 



CONCLUSIONS 



An investigation of the effect of delay 
time on fragmentation was conducted at 
a reduced scale using three blastholes 
per shot in a 45-in bench. With the bur- 
den constant at 15 in, delay intervals 
were varied from 0.0 (simultaneous) to 
36.0 ms/ft of burden, and the entire 
muckpile was screened to assess frag- 
mentation for tests with 21- and 30-in 
blasthole spacings. 

Attempts were made to mathematically 
describe the distribution of fragment 
sizes using power and Weibull functions, 
which have been used by other research- 
ers. However, regression analysis indi- 
cated the Bureau data were best described 
by a Gaussian or simple normal distribu- 
tion. Materials excluded from the analy- 
sis were overbreak, which came from out- 
side the shot area, and fines, which were 



determined to come from the immediate 
blasthole vicinity and showed little var- 
iability from shot to shot. 

Analysis-of-variance tests showed that 
delay time did influence the distribution 
of fragment sizes for both spacings, but 
more so for the 21-in spacing. Fragmen- 
tation was coarsest for shots fired 
simultaneously and at delay times of 
24 ms/ft of burden and greater. Better 
fragmentation was observed for delay 
times from 1 to 17 ms/ft, with the tests 
at 11.2 and 16.8 ms/ft resulting in the 
best fragmentation. Dynamic strain and 
pressure measurements indicated that this 
improved fragmentation may be the result 
of strains induced by stress waves con- 
structively interacting with strains in- 
duced by gas pressure from an earlier 
detonated hole. 



REFERENCES 



1. Chiappetta, R. F. , S. L. Burchell, 
D. A. Anderson, and J. W. Reil. Effect 
of Precise Delay Times on Blasting Pro- 
ductivity, Ground Vibrations, Airblast, 
Energy Consumption and Oversize. Paper 
in Proceedings of the 12th Annual Con- 
ference on Explosives and Blasting Tech- 
nique, ed. by C. J. Konya (Atlanta, GA, 
Feb. 1986). Soc. Explos. Eng. , Mont- 
ville, OH, 1986, pp. 213-240. 

2. Reil, J. W. , D. A. Anderson, A. P. 
Ritter, D. A. Clark, S. R. Winzer, and 
A. J. Petro. Geologic Factors Affect- 
ing Vibration From Surface Mine Blast- 
ing (contract H0222009, Vibra-Tech Eng., 
Inc). BuMines OFR 33-86, 1985, 204 pp.; 
NTIS PB 86-175858. 

3. Bleakney, E. L. A Study of Frag- 
mentation and Ground Vibration With Air 
Space in the Blasthole. M.S. Thesis, 
Univ. MO-Rolla, 1984, 92 pp. 

4. Brinkraan, J. R. The Influence of 
Explosive Primer Location on Fragmenta- 
tion and Ground Vibrations for Bench 
Blasts in Dolomitic Rock. M.S. Thesis, 
Univ. MO-Rolla, 1982, 83 pp. 



5. Smith, N. S. Burden Rock Stiffness 
and Its Effect on Fragmentation in Bench 
Blasting. Ph.D. Thesis, Univ. MO-Rolla, 
1976, 148 pp. 

6. Reed, R. P. Triaxial Measurement 
of Stress Waves in the Free-Field. San- 
dia Lab., Albuquerque, NM, Rep. SAND-78- 
212236, 1979, 21 pp. 

7. Anderson, D. A., S. R. Winzer, and 
A. P. Ritter. Time-Histories of Princi- 
pal Strains Generated in Rock by Cylin- 
drical Explosive Charges. Paper in Rock 
Mechanics in Productivity and Protection, 
ed. by C. H. Dowding and M. M. Singh 
(Proc. 25th U.S. Symp. on Rock Mechanics, 
Evanston, IL, June 25-27, 1984). Soc. 
Min. Eng. AIME, 1984, pp. 959-968. 

8. Da Gama, D. Use of Comminution 
Theory To Predict Fragmentation of 
Jointed Rock Masses Subjected to Blast- 
ing. Paper in First International Sympo- 
sium on Rock Fragmentation by Blasting. 
Lulea Univ. Technol. , Lulea, Sweden, 
1983, pp. 565-579. 

9. Cunningham, C. The Kuz-Ram Model 
for Prediction of Fragmentation From 



95 



Blasting. Paper in First International 
Symposium on Rock Fragmentation by Blast- 
ing. Lulea Univ. Technol. , Lulea, Swe- 
den, 1983, pp. 439-453. 

10. Hjelmberg, H. Some Ideas on How 
To Improve Calculations of the Fragment 
Size Distribution in Bench Blasting. 
Paper in First International Symposium on 
Rock Fragmentation by Blasting. Lulea 
Univ. Technol. , Lulea, Sweden, 1983, 
pp. 469-494. 

11. Siskind, D. E., R. C. Steckley, 
and J. J. Olson. Fracturing in the Zone 
Around a Blasthole, White Pine, Mich. 
BuMines RI 7753, 1973, 20 pp. 

12. Siskind, D. E., and R. R. Fumanti. 
Blast-Produced Fractures in Lithonia 
Granite. BuMines RI 7901, 1974, 38 pp. 

13. Olson, J. J., R. J. Williard, 
D. E. Fogelson, and K. E. Hjelmstad. 



Rock Damage From Small Charge Blasting in 
Granite. BuMines RI 7751, 1973, 44 pp. 

14. Bergmann, 0. R. , F. C. Wu, and 
J. W. Edl. Model Rock Blasting Measures 
Effect of Delays and Hole Patterns on 
Rock Fragmentation. Eng. and Min. J. , v. 
175, June 1974, pp. 124-127. 

15. Winzer, S. R. , D. A. Anderson, and 
A. P. Ritter. Rock Fragmentation by Ex- 
plosives. Paper in First International 
Symposium on Rock Fragmentation by Blast- 
ing. Lulea Univ. Technol., Lulea, Swe- 
den, 1983, pp. 225-249. 

16. Holloway, D. C, D. B. Barker, and 
W. L. Fourney. Dynamic Crack Propagation 
in Rock Plates. Paper in the State of 
the Art in Rock Mechanics (Proc. 21st 
U.S. Symp. on Rock Mechanics, Rolla, MO, 
May 28-30, 1980). Univ. MO-Rolla, 1980, 
pp. 371-379. 



96 



BLASTING EFFECTS ON APPALACHIAN WATER WELLS 
By David E. Siskind 1 and John W. Kopp 2 



ABSTRACT 



The Bureau of Mines, in a contract 
study, examined blasting vibration im- 
pacts on low-yield domestic water wells 
in the Appalachian coal mining region. 
Researchers surveyed 36 case histories to 
determine if blasting was likely to have 
caused the claimed or observed changes, 
ranging from turbidity to loss of water. 
Following these investigations, they con- 
ducted field studies at four sites where 
the impacts of surface mine blasting 
could be directly measured on operating 
wells of known capacities. 

Researchers found no evidence of blast- 
ing effects at the 36 well sites; 
instead, they observed other more likely 



causes. In the field tests, research- 
ers found no significant direct effects 
from the blasting. However, in three of 
the four cases, they did observe changes 
in the static water levels and specific 
well capacities as the excavations ap- 
proached to within 300 ft. Researchers 
attributed these changes to mass rock 
movement resulting from downslope lateral 
stress relief in the low-yield fracture 
system aquifers. With sufficient re- 
charge, static levels recovered and ca- 
pacities increased, provided that the 
nearby mine excavations did not drain the 
aquifers. 



INTRODUCTION 



At about the time the Bureau of Mines 
was studying the problems of dynamic vi- 
bration response and safe levels for 
houses near surface mine blasting, al- 
legations were being made that resi- 
dential water wells were also being 
damaged by blasting. Technical experts 
believed that such effects were unlikely. 
However, there had never been a carefully 
designed and controlled study of this 
problem. Such a study appeared justified 
by the number of alleged cases, particu- 
larly in the Appalachia coal mining 
region. 

The Bureau contracted with Philip R. 
Berger and Associates, Inc., to examine 
the problem of possible vibration damage 
to residential water wells from nearby 
surface mine blasting. The Berger team, 



headed by Donelson Robertson, reported 
their research in a series of three con- 
tract final reports available for inspec- 
tion at Bureau centers and for pur- 
chase through the National Technical 
Information Service (1_~2_)»^ This paper 
summarizes the key findings, which were 
published in November 1980 as volume 1 of 
the contract final report (_1_). 

The study consisted of three parts: 
(1) a background review of vibration and 
other impacts on water wells, such as 
earthquakes, earth tides, and nuclear 
blasts, (2) examination of 36 cases of 
alleged damage from blasting in Ap- 
palachia, and (3) a careful study of 
blasting effects on wells at four surface 
mine sites in Appalachia. 



GENERAL REVIEW OF VIBRATION EFFECTS ON WELLS 



The background review found little that 
was directly applicable. Observed cases 
of well damage were caused by permanent 

Supervisory geophysicis t. 
^Mining engineer. 
Twin Cities Research Center, Bureau of 
Mines, Minneapolis, MM. 



ground displacement, such as land slid- 
ing, rather than vibration. The types of 
effects observed required vibration 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



levels many orders of magnitude higher 
than typical blasting vibrations and were 
listed as "casing collapse, earth dis- 
placement, pump base displacement, mis- 
alignment of pump column," etc. Cases 



97 



specifically involving raining were con- 
cerned with pit excavations and included 
interception with the aquifer, pumping 
from bit bottoms, and ground water 
pollution. 



INVESTIGATION OF REPORTED CASES OF ALLEGED BLASTING 
DAMAGE TO WELLS IN APPALACHIA 



Inquiries to Appalachian surface mines, 
regulatory agencies, explosives sup- 
pliers, coal companies, insurance compa- 
nies, and trade associations identified 
36 reports of blast damage to wells. Of 
these, 24 sites were visited for either 
direct well measurements or discussions 
with owners and/or mine operators. 

In the Berger report, Robertson states: 

In many cases, it was apparent 
that the damage claimed was caused 
by something other than blasting. 
In other cases, it was clear that 
there had been a general lowering 
of the water table, possibly as the 
result of unplugged flowing test 
holes, drainage at the high wall, 
or a two-to-threefold increase in 
the number of residences utilizing 
a limited supply, combined with 
seasonal changes. 

In nearly every case, there was a 
lack of good benchmark data. Many 
residents have only a vague Idea of 
the depth of their wells. Fewer 
know the depth of the casing. None 
of the residents interviewed knew 
the source of the water in their 
wells. About 50 pet had a vague 
idea of the static water level in 
the well when it was initially com- 
pleted. Only one well had been 



tested in any quantitative way. 
That test was inadequate and made 
the owner think he had a much bet- 
ter well than was actually the 
case. 

Consequently, it was very diffi- 
cult to confirm or deny that blast 
damage had occurred, but among the 
36 examples, some of the well his- 
tories suggested two scenarios in 
which blasting might cause damage. 
The first is that the ground vibra- 
tions might be sufficient at times 
to cause loose material such as 
drill cuttings to slough off the 
uncased borehole and cause the 
water to become temporarily turbid, 
or if enough material was involved, 
to bury pump components at the bot- 
tom of the well. The second con- 
cerns those wells that obtain their 
water from flooded and abandoned 
deep mine workings. Ground vibra- 
tions might be sufficient at times 
to cause roof falls that could stir 
up sediment in the water or disturb 
an existing potable water-mine acid 
stratification. Of course, slough- 
ing of the well bore and mine roof 
falls can occur in the absence of 
blasting, so these scenarios are 
not exclusive. 



APPALACIAN WATER WELLS 



GROUND WATER OCCURRENCE IN APPALACHIA 

Most ground water used for domestic 
supplies in Appalachian coal-bearing 
strata are in vertical fractures, joints, 
and along bedding planes. Some of these 
joints are tectonic in origin and have a 
regional pattern. However, local frac- 
ture systems exist from lateral stress 



relief associated with natural topo- 
graphic development. The coal seams of- 
ten serve as the primary water conduit, 
being low in tensile strength and having 
extensive vertical fractures. Often, the 
coal is underlain by relatively fracture- 
free claylike rock preventing further 
vertical migration. In their study, Ber- 
ger engineers found that local systems 



98 



TABLE 1. - Appalachian well water characteristics 

Iron Commonly exceeded recommended 0.3 mg/L. 

Manganese Often exceeded standard of 0.05 mg/L. 

Sulfates 14 to 240 mg/L, below recommended level of 250. 

Total solids: 

Suspended Within acceptable ranges. 

Dissolved Do. 

pH levels Most were within 6 to 8, with the total range 

of 5 to 8.7. 

Color Within acceptable ranges. 

Odor Do. 

Turbidity Commonly exceeded standard limit of 5 units. 



did not always interact, with static wa- 
ter levels sometimes varying between 
wells only 10 to 35 ft apart. 

WATER CHARACTER AND QUALITY 

Information on water quality was ob- 
tained from the literature and tests made 
on wells for this project. The general 
results are summarized in table 1. The 
only observed problem was an occasional 
instance of rusty or reddish-colored 
water seemingly unrelated to the specific 
iron content. In other studies, this 
coloration (red slime) was found to cor- 
respond to the presence of iron bacteria 
and was a problem when wells went un- 
pumped for a long time. The sulfate 
levels suggest no serious acid mine 
drainage nor influence from acid rain, 
which is typically of pH 4.5 to 5 in the 
region. 

TYPICAL DOMESTIC WELLS IN APPALACHIA 

Residential wells in Appalachia are 
typically 6-in-diameter rotary or cable- 
tool drilled and 100 to 150 ft deep 
(maximum about 400 ft). Normally, only 
the top 20 ft or so is cased. The re- 
mainder is unprotected from sloughing off 
the well sidewalls, which is a common and 
normal occurrence. (Recently, it has 



become popular to line all wells deeper 
than 100 ft to control sloughing prob- 
lems.) Important considerations that to- 
gether determine the continuous capacity 
of the well are the pump depth below the 
static water table, storage capacity of 
the well, natural recharge rate, and pump 
size. 

The response to pumping these low-yield 
fracture water-table systems is a rapid 
drawdown until a near-equilibrium situa- 
tion is reached. The pump must be suf- 
ficiently below the static water table 
(submergence) to allow this drawdown and 
still retain water above it. Larger 
pumps produce greater drawdown and re- 
quire correspondingly deeper submergence, 
unless the flow is restricted by a valve 
arrangement. Additionally, rapid draw- 
downs from high pumping rates can cause 
abrupt pressure changes, turbid water, 
and possible "sanding up" of the pump. 

As an example, Robertson calculates 
that an increase of pump submergence from 
50 to 100 ft allows a constant pumping 
rate from a well with only one-fourth the 
specific capacity. He states that a 
pumping rate of 5 gal/min could be 
maintained by the typical Appalachia 
fracture-system well (specific capacity 
of 0.093 (gal/min)/ft of drawdown) pro- 
vided it has 100 ft of pump submergence. 



EXPERIMENTAL STUDY OF VIBRATION IMPACTS ON WATER WELLS 



EXPERIMENTAL DESIGN 

Researchers conducted an evaluation of 
four producing wells to specifically 
determine the influence of blasting on 
well productive capacity. At four sites, 



shallow wells were drilled to obtain wa- 
ter from the coal measures. Additional- 
ly, deep wells drew water from the layers 
below the coal and were isolated from the 
upper measures. For each of these test 
wells, two or three observation wells 



99 



were used to monitor drawdown at varying 
distances. 

Dynamic effects were measured by 10-h 
drawdown tests before, during, and fol- 
lowing blasting. The static water level 
was monitored with float gauges. Water 
quality was also sampled periodically. 
To monitor blasting vibrations, research- 
ers placed seismic instrumentation on the 
surface and at the bottom of an observa- 
tion well for the 1-yr test duration. 

TEST SITES 

Four sites were found to be suitable 
for the tests, allowing sufficient moni- 
toring timt and standoff distance, and 
within convenient travel distance from 
Pittsburgh, PA, where Philip R. Berger 
and Associates is headquartered (fig. 1). 
A summary of the four test sites and 




FIGURE 1.— Location of test sites. 

experimental parameters is given in table 
2. 



RESULTS 



Essentially, the same study was done at 
the four sites in Appalachia. A capsule 
summary is given in table 3. 

BROTHERTON SITE 

Researchers observed a correlation be- 
tween the raining and well changes. The 
closest blast removed supporting material 
(e.g., toe), allowing lateral stress re- 
lief and the widening of vertical frac- 
tures. In other words, the pit excava- 
tion and not the vibrations influenced 
the fracture system. Changes were ob- 
served in the shallow well but not in the 
deep well, likely because of the more ex- 
tensive vertical fractures near the sur- 
face. Significantly, the drops in static 
water level corresponded to increased 
specific capacity. The wider cracks al- 
lowed a temporary drop in static level, 
which would recover with sufficient rain- 
fall. However, the cracks also provided 
increased flow (specific capacity). 

Turbidity results were harder to inter- 
pret. Irregular fluctuations occurred 
during the 11-month study, and some tem- 
porary increases could have been from 
blasting. The nature of the tests con- 
tributed to the turbidity problem: in- 
frequent and periodic pumping, suspended 



iron and drill cuttings, and disturbance 
from the manual sample-collecting proce- 
dure. A continuously used well would not 
have some of these problems. 

TENMILE SITE 

Results were similar to those at Bro- 
therton, except for smaller changes in 
static water levels. Specific capacities 
increased during the 12-month test for 
both shallow and deep wells. They went 
from initial values of about 0.065 and 
0.020 (gal/min)/ft to 0.66 and 0.05, re- 
spectively. As at Brotherton, this was 
attributed to removal of downslope rock, 
resulting lateral stress relief, and the 
consequent widening of vertical frac- 
tures. Along with the increased capacity 
was the availability of increased re- 
charge from the coal seam into the tight 
sandstone formation, possibly accounting 
for the little observed change in static 
water levels. Figures 2 through 5 show 
the site plan, profile, well arrangement, 
and a drawdown test record for the date 
June 9, 1979, when an 0. 80-in/s particle 
velocity was recorded. 

Researchers concluded that ground vi- 
brations produced no deleterious effects 
on either the deep or shallow well, with 



100 











CD 


• 












1 








i—i CO 


Tj rH CD 




i— 1 




i— 1 


<u 


CD 








rH 3 


^-- rH a, 




i— 1 


• 


i— 1 


> 


Xi 








CD 


rH CD O 




CD 


bO 


CD 


o 










12 CD P 


CO E$ rH 




r2 


e 


ES 


X 


X) 








> t-l 


bO co 






•H 




CO 


CD X) 








O P-. 








d 






TJ CD 








• X 


o . cu 




• 


•H 


• 


a> 


a X 








CD (0 


O CD CD 




CO 


S 


co 


a. 


CD O 








4-) . 


O 4-> CD 


• 


4-> 




4-> 


o 


CO 




co 




co J-J bO 


« CO 4-> 


bO 


CD 


3 


CD 


rH 


>. CD 




cd 




CD U-l C 


O CD CO 


d 


CD 


o 


CD 


CO 


X) r4 




4-> 




P T-l 


O P 


•H 


4-> 


rH 


P 




d 




O 




O d 


u-n c 


4-> 




a> 




fi 


4-1 CO 




Z 




d «* t-l 


C O 


CO 


d 


X 


d 


o 


C/3 J-) 








3 e 


4-1 3 


03 


3 




3 




CO 








O CD 


CO O CD 


rH 


o 


co 


o 


co 


co 








13 CO r-l 


Tj (0 


rQ 


Tj 


co 


tj 


cO 


. t-{ 








3 3 co 


XI 3 3 




3 


3 


3 


3 


bO X . 








CO t-l 


CD CO 


CD 


CO 




CO 




d co 








P CD 4-J 


CU SH CD 


> 


r-l 


CD 


P 


CD 


•H CD rH 








Tj 4-) T-l 


B Tj P 


O 


Tj 


P 


X) 


4-1 


d rJ rH 








t-l C 


3 t-J 


X 




•H 




•H 


•H O CD 








OO CD -H 


euro co 


CO 


CO 


CO 


<f 


CO 


6 4-1 3 


CD 






< — f 


CM 




l—l 




•—I 






CD 






















c 

tH 


4-1 C •> 




o 


-3- 




<r 




<r 






e 


d O >~. 

CO t-l P 




CM 

• 


• 




• 




00 

• 






c 


P P tH 


CO 


CM 


m 




CM 










CO 


rH CO CJ " 




1 


1 




1 




1 






T-l 


d P O 


d 


<r 


i — i 




<fr 




LO 






X 


CD X i—l 


•H 


o 


.—i 




nO 




CM 






o 


CO t-l (1) 




• 


• 




• 




• 






CO 


Pi > > 




o 
















1— 1 






















CO 






















cu 


n 




















a, 


CD CM 




00 


<r 




O 




co 






< 


TJ CJ\ 

CD d — 




ON 


co 




NO 




LO 

r-l 






u 


rH CO X 




1 


1 




1 




1 






3 


CO P ■— l 




. — 1 


<r 




oo 










o 


y (ox 




• 


• 




• 










LM 


C/5 -r-l 4-) 
X) lm 




CM 


CM 




CM 

• — I 




CM 

CO 






P 






















CO 






















w 


co 




o 


O 




LO 




o 






P 


CD 




o 


OO 




t~^ 




O 






CO 


P CJ 




m 


io 




r-^ 




O 






cd 


CO d P 














« 






0-1 


CO CO M-l 
rH P 




1 


1 




1 




r— 1 
1 






b0 


CQ co 




LO 


o- 




LO 




LO 






d 


t-l 




NO 


nO 




r*- 




CM 






■H 


TJ 




—1 






i — i 




-tf 






a, 






















a 
























3 


















<t 






CU 






d 


nO 


o 




00 




vO 










#t 


•H 


• 


• 




• 




• 






13 




a a; 


s 


<r 


i-H 




NO 










c 


i— 1 


S p 


*-*^ 


I 


1 




1 




1 






CO 


r-l 


3 CO 


rH 




r-. 








•<f- 








CD 


P-i p 


CD 


o 


CM 




oo 




CM 






bO 


3 




bO 


• 


• 




• 




• 






d 








<r 






— i 










•H 


a 






















4-1 






















co 


CD 


*» 




















CO 


CD 


-< X 




















rH 


Q 


rH 4-J 


j-J 


ON 


r-~ 




00 




CO 






CO 




CD CU lm 


nO 


oo 




LO 




NO 










^ CD 




. — I 


•—i 




r-| 




I — 1 






1 




Tj 




















CM 






d 


ON 


r-l 




CD 




NO 






td 




^ 


•rH 


• 


• 




c 




• 






>J 


|—j 


a, cd 


e 


r~- 


<r 




o 










CO 


I—l 


S j-i 


^-», 


1 


1 




z 




1 






<3 


CD 


3 co 


rH 




rH 








-3- 






H 


3 

3 

o 

i— i 


O-i P 


«0 
W 


LO 

• 

CM 


CM 

• 








CM 

• 






























.— I 


M 






















CO 


-H X 




ON 


nO 




CD 




ON 








X 


I— 1 4_> 


P 


O 


-<r 




d 




nO 








CO 


<D a, n-i 


. — i 


— 1 




o 














S CD 










53 














TJ 
















































n 


> 




n 




1 


« 










a 


^ 




4-J 




co 


X 










O 






d 




u 


O 






CD 




4_) 


M 




•H 




•H 








i_) 




P 


CD 




o 




CO 








•r-l 




CD 


rH 




p-i 




■—I 


0) 






co 




X 

4J • 


•H 

a 




CD 


• 


o 


I— 1 

rH 










o < 


d 




CO 


<! 


• 


t-l 










P P-i 


CD 




o 


Cu 


4J 


> 












CQ 


H 




Pi 




C/D 







CD 
CD 
4-J 



4-) 
CD 
CD 
4-) 

U 

d 
o 

LW 

4-) 
CO 

CD 
4-J 
rH 

d 
co 

CD 

u 



>N 

r-l 

S 
3 
en 



W 

rJ 

CQ 
H 





1 CD • 








• 














O CD bO 






CO 


bO P 














a. co d 




l l 


CU 


d <p 












>N 


S cO t-l 




CO tH 


co 


•H 








P 




4-1 


CD CD 4-> 




tH CO 


CO 


P o 


• 






d 




•H 


4-1 V-l CD 




P CD 


CD 


CO O 


d 






CD 




TJ 


cj co 




CO O 


P 


cO co 


o 






tH 




•H 


CD d rH 




> P-. 


O 


rH 


•H 






O 




-O 


rH tH X) 






d 


x d 


p 






tH 




U 


rQ 




>N 


•H 


tH 


(D CO 






LM 


• 


3 


•H >, S 




rH . 




a x 


rH tH 






LM 


cO 


H 


CD U O 




rd CD 


0) 


O P 


P P 






3 


p 




CO CO V-l 




bO rH 


rH 


P tH 


P CO 






CO 


(0 




O r4 lm 




•H XI 


X 


LM 3 


tH > 






d 


TJ 




P-i 




* 






rJ 






M 
















1 


d 




• 




>-. 












tH 1 


CD 


1 


d 




Jj 












P CD 


CD 


O 


o 




XJ 












CO tH 


3 


X 


• tH 




CD 












> TJ 


p 


CO 


CO P 




•H 














CD 


rH 


P CO 




e 












TJ TJ 


X 




CO tH 




CD 


CD 










CD d 




TJ 


CD P 




X 


CD 










d co 


>> 


d 


P CO 




O 


bO 

d 










•H 
CO CO 


a 

d 


CO 


> 




P 


CO 










rH d 


CO 


TJ 


P CD 




CD 


rC 




o 






a, o 


CU rH 


O rH 




4-) 


o 




TJ 






X tH 


CD 


CD 


P P 




cfl 












CD P 


p 


•H 


CO P 




C5 


o 

53 










d to 
d> 


CJ 


LM 


P tH 

rJ 










3 






CD 




• 




1 




TJ 




Tj O 






d 




CD 




tH 




W CD rH 




CD rH 


• 




o 


d 


d 


r^ 


d ; 




O 4-1 rH 




CO rH 


rH 




Tj 


CD 


rH 


rH 


tH 




LM Li CD 




CO CO 


rH 


• 


CD P 


X 


CO 


O 




A 


CO 3 




CD x: 


CD 


NO 


> o 


H 


> 


• 


"^ 


>N 


>N J-> 




P CD 


3 


NO 


O LM 






o 


CD 


4J 


Tj CD 3 




bo 




• 


p 




rH 




d 


tH 4J 


CO O 




O P 


&o 


CL, P 


• 


CO 


o 


•H 


O 14H 


CD d rH 




P O 


CD 




a Oh 


CD 


•H 


p 


CO 


CO ^~- 


4-» CD rH 




a. lm 


CD 


CD 


M CD 


CO 


P 




rH • 


CX-^ 


co x: co 






TJ 


CO 


CJ 


CO 


•H 


r^ 


CU CO 


co d 


4-) rd 




P o 




3 


X 


CD 


d 


O 


X CD 


CJ tH 


CO 




CO — I 


P 




• CD 


P 


•H 


O 


CD 3 


s 


<f TJ N^ 




CD 


O 


CD 


co 


O 




• 


d rH 


o >-, 


co d 




P LM 


LW 


d 


co p 


CD 


O 


o 


3 CO 


•H rH 


• CO CD 




o 




rH 


• CO 


TJ 


P 




> 


LM CO 


O CD 




CO 


co 


CO 


O CD 






p 


P 


•H W) 


CD CO 




cO P 




> 


P 


>nT5 


CO 


O X 


o ^^ 


>>rd <V 




o 


TJ 




>N 


P 


0) 




ip bO 


CD 


rH 4-J P 




TJ P 


d 


P 


rH d 


CO 


P 


3 


•H 


CU 


rH d O 


• 


CD O 


CO 


CO 


rH «iH 


P 


CD 


o 


P X 


C/l 


to o d 


•—\ 


> co 




CD 


CO 


O 


> 


rH 


Cu 




t-l S tH 


>, 


O MH 


rH 


X 


tH CD 


& 


o 




CD <-^ 




4-> 


rH 


P 


rH 


bO 


P P 


a 


o 


>N 


O CO 




•H rH O 


d 


a, >. 


CD 


•H 


•H CO 


CD 


CD 


P 


X tH 




d rH 4-1 


o 


a x 


3 


X 


d rH 


P 


P 


CD 


CD p 




M 




M 






M 






> 






B u 




p 






P 












o o 




O X 






O 






P 






W LM rH 




LP X 


P 




LM 






d 






LM rH 


o 




o 










CD 






CD CD 


p 


On >, <4_: 




P 






a 






CD CX 3 




O P 






a 






a, 






rd O 


TJ 


P CD 


CD 




CD 






tH 




rH 


4-1 u 3 


CD 


TJ > 


bO 




CJ • 






d 




CD 


d tj o 


P 


o 


d 




X CD 






a* 




> 


O rH 


CO 


CD O 


CO 




CD CO 






CD 




CD 


S 4-1 rH 


rH 


a a) 


X 




CO 










i— 1 


-d CO 

— i ho X 


CD 
P 


O P 
CO 


CJ 




d CD 
O P 






e 

o 




t-l 


— 1 t-l CD 




X 


CD 




tH CJ 






p 




CD 


rH n— ' 


>, 


p 


rH 




p d 






LM 




4-1 


d to 


P 


• t-l 


p 




CO tH 










CO 


•H CO 


CD 


d 3 


p 




tH 






CO 




3 


4-1 


> 


o 


•H 




P >N 






p 






4-1 . CD 


o 


tH rH 


rJ 




cO P 






CO 




CJ 


lm bO cO 


o 


P rH 






> CO 






TJ 




•H 


d rH 


CD 


cO CD 




* 


P 










4-) 


— 1 t-l X 


Pi 


. tH 3 


• 


i— i 


P o 






CD 




CO 


cm a. 




rH P 


CD 


rH 


o cu 






P 


• 


4-> 


S 4-1 




rH CO 3 


bO 


CD 


d B 






CD 


CD 


C/l 


TJ d CD 


• 


CO > O 


P 


3 


12 






rH 


P 




CD O, CD 


^-N 


LM y-f 


CO 








& 


3 




<X CD 


>% 


d P rH 


X 


a. 








a 


rH 




CX 4-1 O 


rH 


•H O CO 


o 


CD 


>> CD 






o 


•H 




O -H rH 


d 


CO d X 


CD 


CD 


p d 






CJ 


CO 




u o. o 


o 


P -H CO 


p 


TJ 


CD O 






d 


LM 




Q 




S 






> 






M 






d 




^ 






P 






i 

CO 


• 

IX! 




o 










d 






p 


O 


CD 


4-1 




a 






•r-l 






tH 




4J 


P 




CD 






O 






« 


«« 


•H 


CD 




rH 






P-i 






rH 


CD 


C/l 


X 
P • 

o < 
p Pu 

CQ 




•H 

a 

d 

CD 
H 






CD • 

co <3 
O P-i 
Pi 






CJ) 

• 

CO 


tH 
> 



101 




FIGURE 2.— Tenmile test site plan. 




50 100 

PUMPING TIME, min 



,000 



FIGURE 5.— Tenmile drawdown test for deep well (D-1). 



ROSE POINT SITE 



I.SOO 1 - 



Well 
D-2 



Q 



FIGURE 3.— Tenmile test site profile. 



Well 
D-l 



Well Well 

S-l S-2 



200 



Well 
S-3 



o o o o o 

L — | o ft — »U — i o ft — »l« — 7 ft— J* — 8 ft — >T 

200 ft 200 ft 160 ft 160 ft 160 ft 

10 

Scale, ft 

FIGURE 4. — Tenmile well layout. Deep and shallow pumped 
wells are D-1 and S-1. Others are observation wells. 

one possible exception. A blast that 
went offscale at 2 in/s (distance of 
85 ft) could have dislodged and caused a 
loose rock in the sidewall to shift into 
the hole, producing a partial bridge and 
preventing the sounding of the bottom. 

Turbidity results were also similar to 
those from the Brotherton tests, with 
wide variations, difficult measurement 
conditions, and the suggestion of tem- 
porary increases from close-in blasts, 
less than 300 ft away. 



This site differed from the previous 
two in that the well was below the blast- 
ing. Consequently, researchers expected 
little stress relief effect. Indeed, 
little change occurred in either static 
water level or specific capacity until 5 
months into the test, when specific capa- 
city jumped from about 0.33 to nearly 
0.60 (gal/min)/ft and later decreased. 
Researchers attributed the changes to 
nearby removal and replacement of over- 
burden, which occurred at that time. As 
with previous sites, no direct effect of 
blasting was evident. 

ST. CLAIRSVILLE SITE 

This site was characterized by very low 
capacity wells, which were expected to be 
susceptible to outside disturbances. The 
initial drawdown test in the shallow well 
appeared satisfactory, with the well 
able to sustain pumping rates of 
0.86 gal/min, equivalent to a specific 
capacity of 0.041 (gal/min)/f t. 

The deep well was worse, and was only 
able to maintain 0.21 gal/min for about 
half the normal test period of 600 min. 
(For this site, as well as the others, an 
unperforated liner and packer were used 
to isolate the deeper well from recharge 
from the shallower coal measure being 
mined. ) 

One month after the initial drawdown 
test, the shallow well was again pumped, 



102 



but this time it had a specific capacity 
of only about one-third as much and at 
about one-half the pumping rate. Re- 
searchers suspected that air had been 
entrapped from initial tests, preventing 
full recharge. 

In summary, researchers observed no 
clear blasting effects at this site. Un- 
fortunately, work stopped short of dis- 
tances found to produce effects observed 
at the three other sites. Some equipment 
failures were also experienced. 

DOWN-THE-HOLE VIBRATIONS 

Vibrations were monitored at the bottom 
of shallow observation wells as well as 
on the surface. Downhole vibrations 
were, as expected, of lower amplitudes, 
suggesting less risk to subsurface as 
opposed to surface structures. Table 4 



summarizes the average relative resultant 
velocity amplitudes. 

TABLE 4. - Reduced vibrations 
measured at depth 



Site 



Brotherton, PA.... 

Tenmile, WV 

Rose Point, PA.... 
St. Clairsville, 
OH 



Depth, ft 



149 

160 
168 

180 



Relative 
amplitude 1 



: 0.34 

3 .68 

.44 

.14 

-.25 



1 Depth vibration divided by surface 
vibration. 

2 Blasting in poorly confined upper 
layers. 

3 Blasting in well-confined lower 
layers. 



CONCLUSIONS 



Research at four sites in Appalachia 
found no catastrophic effects on water 
wells from blasting at vibration levels 
up to about 2.0 in/s. At three of the 
sites (and possibly the fourth, had test- 
ing continued), long-terra changes were 
observed and were attributed to the 
removal of confining rock. 

As blasting and the pit excavation ap- 
proached within 300 ft of the wells, the 
mechanism of lateral stress relief al- 
lowed vertical fractures to open. Be- 
cause these fracture systems are typi- 
cally the abode and conduit of shallow 
Appalachian ground water, the static 
water levels then dropped over a period 
of weeks. With sufficient rainfall, the 
water levels would return. Where suf- 
ficient submergence exists, such minor 
changes in static level would not be 



noticed. Of benefit to the well user is 
the increased storage and flow as shown 
by higher observed specific capacities. 
Shallow wells exhibited this effect more 
than deep wells, consistent with expecta- 
tions that more extensive fracture sys- 
tems exist at shallow depths. At one 
site, backfilling reversed the improve- 
ment, either from clogging by fines or 
reintroduction of crack-closing lateral 
confinement. 

Blasting may cause some temporary in- 
creases in turbidity, of the same order 
as those occurring in the absence of 
blasting. Results were uncertain on this 
because of the difficulty of sampling 
without causing disturbance and natural 
sloughing. Plastic well liners were 
recommended to control turbidity. 



REFERENCES 



1. Robertson, D. A., J. A. Gould, J. 
A. Straw, and M. A. Dayton. Survey of 
Blasting Effects on Ground Water Supplies 
in Appalachia (contract J0285029, Philip 
R. Berger and Associates, Inc.). Volume 
I. BuMines 0FR 8(l)-82, 1980, 159 pp.; 
NTIS PB 82-152125. Volume II. BuMines 
OFR 8(2)-82, 1980, 266 pp.; NTIS PB 82- 
152133. 



2. Berger, P. R. , D. T. Froedge, J. A. 
Gould, and L. F. Kreps. Survey of Blast- 
ing Effects on Ground Water Supplies in 
Appalachia. Part II (contract J0285029, 
Philip R. Berger and Associates, Inc.). 
BuMines OFR 188-83, 1982, 114 pp.; NTIS 
PB 84-113182. 



103 



FIBER OPTIC PROBE TO MEASURE DOWNHOLE DETONATION 
VELOCITIES OF EXPLOSIVE COLUMNS 

By David L. Schulz 1 



ABSTRACT 



Following ideas developed by research- 
ers at the University of Maryland, the 
Bureau of Mines assembled a versatile, 
readily available, and very inexpensive 



fiber optic probe for downhole measure- 
ment of explosive shock front position 
and velocity. The accuracy of the probe 
was determined in field testing. 



INTRODUCTION 



Detonation velocities of explosive col- 
umns are often measured to determine in 
situ explosive performance. Several 
methods are in use, ranging from simple 
resistance probes to very sophisticated 
and expensive electromagnetic resonance 
measuring systems. 

In a recent study on blasting, under 
Bureau of Mines contract S0245046, 



University of Maryland researchers used a 
fiber optic system to measure detona- 
tion velocity. Discussions with these 
researchers and further Bureau work led 
to the development of the fiber optic 
probe described in this paper. 



SYSTEM DESCRIPTION 



The fiber optic probe consists of a 
fiber optic cable to detect and carry 
the detonation-zone light emission and 
a sensor to detect and convert the 
light signal into an electrical signal. 
Figure 1 shows the Bureau's system. 

FIBER OPTIC CABLE 

The fiber cable is a DuPont Crofon 2 
lightguide consisting of sixteen 0.010- 
in-diameter optic strands in a plastic 
tube with an overall outside diameter of 
0.087 in. The lightguide used in the 
Bureau's tests was obtained from the 
Edmund Scientific Co. at a cost of about 
$0.70/ft in September 1986. 

LIGHT-DETECTING SENSOR 

The light detector is a fiber optic 
data link receiver module-100 series with 
the trade name Optolink. Designed for 

Electronics technician, Twin Cities 
Research Center, Bureau of Mines, Minne- 
apolis, MN . 

2 Reference to specific products and 
sources does not imply endorsement by the 
Bureau of Mines. 



data transmission applications, the mod- 
ules cost about $16 each and are re- 
usable. Bright light through the system 





FIGURE 1.— Fiber optic detector and cable. Cable outside 
diameter is 0.087 in. 



104 



gives an output of about 100 to 300 mV. 
These particular modules by Interoptics 
were obtained through the Newark Elec- 
tronics catalog. 

The receiver module incorporates a 
photo diode to detect the signals, and a 



receiver-integrated circuit chip to am- 
plify the signals in one package. The 
overall dimensions are approximately 1/2 
by 5/8 by 7/8 in. 



APPLICATION NOTES 



The fiber optic cables are 
length with a sharp tool to get 
light-conducting surface, and one 
inserted into the light detector 
other end of the cable is embedded 
explosive column from the top down 
predetermined distance from the 
tion point, usually the bottom of 
plosive column. Intense light i 
off as the detonation front moves 
explosive column. This light is 
up by the fiber optic as the det 
passes the ends of the fiber 
Carried by the fiber optic cable 



cut to sensor, the light signals are converted 

a good to millivolt signals. Typical rise time 

end is is 0.01 ms. The millivolt signals are 

The measured by an oscilloscope using a raem- 

in the ory or hold feature to retain both the 

to the voltage spikes for each bundle of fibers 

initia- and an initiation-time spike. From known 

the ex- insertion distances and measured times, 

s given velocities are simply calculated. A 

up the four-channel Nicolet digital oscillo- 

picked scope, model 4094A, set to a time scale 

onation of 0.5 us per point, was used for the 

cables. Bureau's tests. 

to the 

FIELD TESTS 



Bureau researchers used the fiber optic 
system for two series of field studies of 
explosively produced rock fragmentation. 
The first tests used explosive columns of 
7/16-in-inside-diameter, 40-in-long plas- 
tic tubes filled with 60-pct-extra dyna- 
mite. The measured detonation velocity 
was 8,680 ft/s at the bottom of the col- 
umn where the dynamite was densely 
packed. At the top where the powder was 
less well packed, the measured velocity 
was 7,350 ft/s (fig. 2). In 5/8- 
in-inside-diaraeter columns, measured ve- 
locities were 9,920 to 11,400 ft/s. 



A second set of tests was run on larger 
charge diameters. Explosive columns were 
16 ft long and consisted of 2-in sticks 
of 60-pct-extra dynamite. Holes for the 
fiber optic bundles were punched in the 
sticks at desired distances from the ini- 
tiation point. The optic cables were 
then inserted and taped in place and the 
powder cartridges carefully loaded. Ac- 
curacy was determined by the dimensional 
stability of the column of powder car- 
tridges (some compaction can occur). Re- 
searchers measured a steady-state ve- 
locity of 15,140 ft/s. 



105 



Probe B 



Probe A 



,F 



20' 



i 



I I 

-M 




366 /is 



TIME, /*s 
FIGURE 2.— Outputs from fiber optic probe for bottom-initiated blasthole showing measured velocity between probes of 7,353 ft/s. 

CONCLUSIONS 



Advances in fiber optic technology have 
provided an inexpensive and versatile 
method to measure explosive detonation 
velocity. An off-the-shelf system can be 
assembled and accurate mesurements made 



of detonation velocity for as little as 
$10 per blast. Only a memory oscillo- 
scope is needed, in addition to the fiber 
optic cable and detector module. 



106 



STEMMING EJECTION AND BURDEN MOVEMENTS 
OF SMALL BOREHOLE BLASTS 



by John W. Kopp 1 



ABSTRACT 



Stemming is used in blasting operations 
to help contain explosive gases as long 
as possible. Stemming can reduce air- 
blast, improve fragmentation, and reduce 
the chances of hot explosive gases ignit- 
ing methane and dust explosions in under- 
ground mines. 

The types and amounts of stemming ma- 
terial that are desirable in underground 
metal and nonmetal mine blasting to im- 
prove fragmentation while containing the 
hot gases are largely unknown. This Bu- 
reau of Mines research examined the ef- 
fectiveness of differing lengths of stem- 
ming by measuring stemming ejection times 
as related to burden movement. With 



properly stemmed blasts, stemming is con- 
tained until some burden movement has oc- 
curred. Test blasts at two surface lime- 
stone quarries were evaluated using 
high-speed photography. For the condi- 
tions of these tests, a stemming length 
of at least 26 charge diameters was found 
to prevent premature stemming ejection. 
In tests with stemming lengths of 16 
charge diameters, the stemming was ef- 
fective but there was early venting of 
hot gases through fractures in the rock. 
Further testing with other rock types, 
hole diameters, explosive types, and 
stemming materials to determine their 
effect on incendivity is recommended. 



INTRODUCTION 



Methane emissions in underground mines 
can present hazards, especially when ig- 
nition sources such as hot gases from 
explosives are present. The problems 
associated with blasting in underground 
coal mines have been addressed by use of 
permissible explosives and permissible 
procedures for their use. However, meth- 
ane also occurs in some noncoal under- 
ground mines, particularly oil shale, 
trona, salt, potash, copper, limestone, 
and uranium mines. At present, such 
mines receive a variance from the U.S. 
Mine Safety and Health Administration 
(MSHA) blasting regulations, depending on 
the source of methane, associated ore 
body, and the method of mining. Conven- 
tional explosives and blasting agents, 
rather than permissible explosives, are 
normally used for both practical and 
economic reasons. 

A recent Bureau of Mines contract^ 
examined blasting practices in gassy 

'Mining engineer, Twin Cities Research 
Center, Bureau of Mines, Minneapolis, MN . 

Contract JO215031; Bauer, Calder & 
Workman, Inc. 



noncoal mines. Most of these operations 
use conventional explosives in standard 
underground blasting practices. Safety 
is sometimes ensured by evacuating all 
personnel to the surface during the 
blast. However, this is often not prac- 
tical for large mines utilizing mining 
methods such as room and pillar. Some 
mines require 20 or more blasts per day, 
involving large amounts of explosives. 
In order to maintain production, blasts 
must be scheduled while personnel are 
working in the mine. 

The contractor made a number of recom- 
mendations for blasting underground with 
personnel present in the mine. Important 
among these was use of stemming to con- 
tain the hot gases and flame from the ex- 
plosion in the borehole until expansion 
of the burden sufficiently cooled the 
gases to prevent ignition of methane. 
The contractor made some predictions of 
stemming behavior, based on a simple 
mathematical model, and recommended that 
his calculations be confirmed by field 
studies. 

This Bureau study was conducted as 
a followup to the previous study, to 



107 



measure the retention time of various 
lengths of stemming in the borehole dur- 
ing normal blasting and to relate stem- 
ming retention time to the burden move- 
ment caused by the expanding gases of the 
explosive. Tests were conducted at two 
surface limestone quarries, which allowed 



careful control of the test blast design 
variables and adequate lighting for high- 
speed photography. While this study is 
directed at blasting underground, the 
efficient use of stemming will improve 
blasting in surface mines also. 



EXPERIMENTAL DESIGN AND PROCEDURE 



Early experimentation had shown that 
high-speed cinematography was the best 
method for measuring stemming movement. 
This method also allowed observation of 
the burden displacement and dust and 
smoke generated by the blast. 

Field experiments were filmed with two 
cameras, a 16-mm rotating-prism camera 
capable of speeds up to 11,000 frames per 
second and a 16-mm registering-pin camera 
with filming speeds up to 500 frames per 
second. The registering-pin camera has 
better resolution than the rotating-prism 
camera and thus provides a much clearer 
picture. 

The time of detonation of the explosive 
was recorded on film with Nonel^ shock 
tubing. A known length of shock tubing 
was attached to the explosive charge and 
passed through the stemming to the sur- 
face and coiled to allow the flash to be 
recorded on film. Detonation of the ex- 
plosive initiated the tubing, which deto- 
nated at 6,000 ft/s. Thus, the time of 
detonation was determined by noting the 
flash of the coiled tubing on the film 
and calculating the time required for the 
detonation to reach the surface. 

Full-scale field tests were performed 
at a surface limestone quarry in order to 
eliminate lighting problems when filming 
with the high-speed cameras. Twelve 

Reference to specific products does 
not imply endorsement by the Bureau of 
Mines . 



cratering shots were detonated in a fac- 
torial experiment to test two types of 
stemming material at three lengths of 
stemming and with two explosive types. 

A high-energy explosive and a rela- 
tively low energy explosive were used for 
this series of tests. The low-energy ex- 
plosive was chosen to produce results 
similar to those from ammonium nitrate 
and fuel oil (ANFO). Both explosives 
were in 1-1/4-in-diameter cartridges. 
Enough explosive was used in each test to 
make a charge 16 in long. The volume of 
explosive was not varied for this test 
series. Properties of the explosives 
used are shown in table 1. 

The blastholes had a 1-1/2-in diameter 
and ranged from 36 to 72 in deep. The 
holes were drilled vertically in a lime- 
stone quarry floor. This represents the 
worst case for blasting efficiency, al- 
lowing relief in only 
upward. 

The stemming material 
crushed limestone in one 
The material was either 
screened to less than 
series mesh size (0.0661 in) or limestone 
gravel between 3/8- and 3/16-in size. 
The stemming material was added above the 
explosive and filled the hole to the col- 
lar. Table 2 shows stemming length and 
type, and assignment of shot numbers. 
The stemming was lightly tamped and had a 
density of about 1.5 g/cm 5 . 



one direction, 

consisted of 

of two sizes: 

drill cuttings 

minus 10 Tyler 



TABLE 1. - Properties of explosives used in the test series 



Explosive 
type 



Density, 
g/cm 3 



Detonation 
velocity, 
ft/s 



Relative 

bulk 
strength 



Explosive 
temperature, 
K 



Borehole 
pressure, 
atm 



High energy 
Low energy 
'ANFO = 100. 



1.16 
1.07 



15,000 
10,500 



148 
115 



3,000 
2,870 



30,000 
19,000 



108 



TABLE 2. - Experimental design and assignment 
of test numbers 





20 


32 


50-60 


Fine drill cuttings: 


S-3 
S-6 

S-4 
S-5 


S-2 
S-12 

S-8 
S-7 


S-ll 




S-l 


Coarse crushed stone: 


S-10 




S-9 







RESULTS 



Twelve cratering 
according to the 
table 2. All sho 
the two high-speed 
viously, one rutin 
second and the o 
frames per second, 
showed that sterami 
from the shallow 
lengths of 20 in, 
was ejected as fol 



Shot 



tests 

factor 

ts were 

cameras 

ing at 5 

ther at 

Analys 
ng was 
holes, 
the st 
lows : 



were conducted 

ial design in 

monitored with 

described pre- 

00 frames per 

1,000 to 3,000 

is of the films 

usually ejected 

With stemming 

eraming material 



Ejection; 
ms 



S-3 13 

S-6 32 

S-4 8.8 

S-5 Retained 

Longer stemming lengths resulted in re- 
tention of the stemming. When stemming 
is retained, it has done its job in terms 
of confining the hot explosive gases- 
However, when stemming is ejected, fur- 
ther analysis is required to determine if 
an adequate length of stemming was used. 

Further analysis of the films indi- 
cated the motion of burden. Not only can 
velocities of various parts of the burden 
be calculated, but an estimate of the in- 
creased burden volume caused by the ex- 
panding gases can be made. Figure 1 
shows stemming and burden movement for 
shot S-3, which was a test with 20 in of 
fine stemming, using the high-energy ex- 
plosive, which resulted in a stemming 
ejection time of 13 ms . Initial movement 
of the stemming was obscured by dust 
caused by venting through cracks in the 
burden. The initial velocity of the 
stemming was calculated to be 190 ft/s. 



The burden velocity was calculated to be 
27 ft/s. The increase in burden volume 
was calculated by assuming the burden 
movement to be in the shape of a cone and 
measuring this increase on figure 1. It 
is apparent from the figure that the bur- 
den movement is closely approximated by a 
cone. The rate of volume increase was 
found to be 600,000 in 5 /s. A plot of 
stemming movement and burden expansion 
versus time elapse from initiation of 
detonation is shown in figure 2 for shot 
S-3. It is apparent from figure 2 that 
considerable burden movement occurred be- 
fore the stemming was completely ejected 
after 13 ms. In this case, some cooling 
of the hot explosive gases had occurred 
because of volume expansion as the gases 
worked their way into the fractured bur- 
den region. 

An estimate of the amount of cooling of 
the explosive gases can be obtained by 
assuming the thermodynamics of the expan- 
sion to be adiabatic. The expansion is 
rapid and allows little time for heat to 
be exchanged between the gases and the 
surrounding rock. From the first law of 
thermodynamics it can be shown that tem- 
perature and volume of the gas are 
related as follows: 



(1) 



T, T_1 V, = T 2 T_1 V 2 ' 



which on rearranging becomes 



= GT' 



1 » 



(2) 



where T 1f T 2 are the initial and final 
temperatures, V 1} V 2 are the initial and 



109 



Burden 

Dust 

Stemming 

4,6 ; etc. Film frame number 







FIGURE 1.— Stemming and burden movements of shot S-3. 



final volumes, and T is the ratio of the 
heat capacities of the expanding gases. 
The actual value of T depends on the 
molecular structure of the gases in- 
volved. Most of the gas products of the 
explosives used are diatomic and poly- 
atomic gases: nitrogen, carbon dioxide, 
and water vapor. The average value of T 
for these gases is approximately 1.3. 

The equation describing the gas temper- 
ature line in figure 2 becomes 



<¥Y- 



i > 



(3) 



where Vj, T\ are the initial volume and 
temperature of the explosive at detona- 
tion, V is the increase in volume, and T 
is the temperature at that volume. 

It is now assumed that the gas expands 
into all of the new volume created by the 
expanding burden. Though the expanding 
gases may not actually fill all of the 



new spaces created by fracturing of the 
burden, this is approximately balanced, 
as no allowance is made for expansion of 
the gases into existing voids or for 
porosity of the rock. The estimated tem- 
perature is thus an approximation but 
provides some insight into the phenomena 
involved. 

Figure 2 also shows the predicted gas 
temperature decrease based on the use of 
equation 3 and the burden volume increase 
as determined from analysis of high-speed 
films. The stemming remained in the 
borehole for 13 ms , at which time the gas 
temperature is estimated to have cooled 
from the detonation temperature of 3,000 
K (2,727° C) to 550 K (277° C), which 
would be sufficient to prevent ignition 
of a methane-air mixture, since the ig- 
nition temperature of methane is 905 K. 
However, from figure 1, it is evident 
that venting, probably through a major 
fracture, occurred before expulsion of 



110 



30 



111 

a: 


c 




^ 








2 


20 


<r 


UJ 




UJ 


5 




a. 


UJ 




5 


> 




UJ 


o 




H 


2 




en 






< 


o 




o 


2 




_i 


2 




< 


2 




TIC 


_ UJ 

H 


10 


UJ 






or 






o 






UJ 






X 







II 1 1 

1 

1 


1 1 1 1 S 1 

X / 


■ 


/ * 


1 KEY 


X s 
X s 


1 


X ' 


\ Burden 


X s 
X s 
X s 
X * 


'. — — Temperature 


X s 


y / 

X ' 


X ' — 


\ 


/ 


\ 


X ' 
X ' 
X ' 
X ' 


\ 


x * 
x * 
X ' 


\ 


X ' 
X ' 

x * 


\ y 


X ' 


\ / 
\ y 


* 
* 
* 

s 


\ / s 


s 


^ / y 




^^x x 




•^^ • 








Ventina / *' 






/ ,'' 






y * 






yy 


Stemming ejected 




yy 






^*&\ 


i i i 


1 1 - 


1 1 1 



12 



8 



14 



16 



18 



10 12 

TIME, ms 
FIGURE 2.— Relative movements of stemming and burden for shot S-3 and the associated gas temperature. 



c 

O 

8 uf 
< 

Ul 

tr 
<_> 

2 

UJ 

_l 
o 
> 

4 2 
UJ 

Q 

a: 

3 
CD 



20 



the stemming, at 3.8 ms after initiation. 
Figure 2 shows that the explosive gases 
would have cooled only to 1,300 K after 
3.8 ms , not sufficient to prevent methane 
ignition. 

A similar analysis was performed for 
shots S-4 and S-6. Results are presented 
in figures 3 and 4. Figure 3 presents 
the analysis of shot S-4 where 20 in of 
coarse stemming was ejected using the 
high-energy explosive. At the time of 
stemming expulsion at 8.8 ms , the explo- 
sive gases were estimated to have cooled 
to 650 K, within a safe temperature 
range. However, venting of dust and 
smoke was observed at 3.5 ms after initi- 
ation. From figure 3, the estimated tem- 
perature of the gases would be 980 K, 
above the safe limit. Again, the venting 
probably occurred through existing frac- 
tures in the rock. This shot and the 
previously discussed shot used a high- 
energy explosive, but the previous shot 
used a different stemming size. The 
finer material held longer. 



Shot S-6 had 20 in of fine stemming and 
the lower volume-energy explosive. Stem- 
ming was ejected from shot S-6, though 
at a slower rate than with the higher 
energy explosive (fig. 4). The time of 
ejection was 32 ms, and the estimated gas 
temperature at the time was 625 K. Vent- 
ing of smoke or dust was also observed, 
starting at 4.8 ms after initiation. 
From figure 4, the estimated gas temper- 
ature at the beginning of venting was 
about 1,300 K, high enough for a methane 
ignition. 

Four shots were fired with 32 in of 
stemming in each hole. The stemming 
remained intact for all these shots. The 
film analysis did not show any stemming 
movement. Burden movement was slower 
than in the previous shots using less 
stemming. With the exception of shot S- 
12, no venting of smoke or dust occurred. 
Also, a rubble zone of broken material 
was left at the surface of each hole. 
The zone was about 2 ft in diameter for 
three of these shots, smaller than the 



Ill 



60 



50 



- ■_" 40 



30 



20 



*: 






O 




„ 




UJ 




a: 


c 


Z> _ 




I- 2 


— t- 


< 


Z 


tr 


UJ 


UJ 


2 


a. 


ui 


5 


> 


LlI 


o 


1- 


5 


co 




< 


o 


o 


z 


_i 


2 


< , 


? 


o 1 


- Ul 


i- 


H 


Ul 


CO 


tr 




o 




UJ 




X 





10 - 



Burden 
Stemming 



Temperature 




2 4 6 8 10 12 14 16 18 

TIME, ms 

FIGURE 3.— Relative movements of stemming and burden for shot S-4 and the associated gas temperature. 



ui 
a: 

=> 
< 

<r 
ui 
o. 

ui 



co 
< 
o 

_) 

1:1 I 

h- 

Ul 

rr 

o 

UI 

X 



ou 




1 1 1 1 1 / 1 ' 






• 




1 


/ 




1 


KEY /' 


25 


1 


Burden / 


— i 


oiemming / 




1 


Temperature / 




1 


/ 


c 
H 20 


- 1 


/ 
/ 


z 

UI 


\ 


^--~~'~ 


2 


\ 


/ .x^ 


UI 

> 


/ . — 


lis 

2 


- \ 


• >^ 


2 




^ / / 


2 




V / x 


y io 





\ / / 


i- 

CO 




/ x 


5 


— 


/ / 




Venting 


/ / 




/ 
/ 


/ Stemming ejected 






/ 


/ 






>' - 


I / I I i i 1 \ 


' 



15 20 

TIME, ms 



25 



30 



O 

uj" 
2 co 
< 
ui 
or 
o 



ui 



o 
> 



UI 

o 
or 

z> 



35 



FIGURE 4.— Relative movements of stemming and burden for shot S-6 and the associated gas temperature. 



112 



2 to 5 ft for 20-in stemmed blasts. The 
presence of rubble zones at the surface 
for both the 20- and 32-in stemming cases 
indicates that with 32 in of stemming, 
the explosive is still sufficiently close 
to the surface to allow fragmentation of 
the burden. 

The final four shots of this series 
used stemming lengths of 50 to 60 in. No 
stemming movement was detected for any of 
these shots. Burden movements were also 
smaller than for the previous shots. 
Figure 5 shows a comparison between bur- 
den velocities and the length of stemming 
used in each hole. The type of explosive 
used made a difference in burden velocity 
only at the two shortest stemming 
lengths. Differences caused by stemming 
type were inconclusive. 

In this crater test series with 1-1/4- 
in charges in 1-1/2-in- diameter blast- 
holes, it was found that stemming was 

30 



KEY 

d High-energy explosive 
O Low-energy explosive 




20 



30 40 50 

STEMMING LENGTH, in 



60 



retained in all cases with stemming re- 
gions of 32 in or greater. When 20 in of 
stemming was used, there was stemming 
ejection in three of the tests after 8.8 
ms or more, but venting of gases through 
fractures occurred before 8.8 ms. Esti- 
mates were made of explosive gas cooling 
associated with volume increases as the 
explosive gases expanded into the frac- 
tured burden. In the three tests where 
stemming ejection occurred, explosive 
gases were estimated to have cooled suf- 
ficiently to prevent ignition of methane 
in the time required for stemming ejec- 
tion (8.8 ms) but not at the time when 
venting of gases through fractures occur- 
red. There was no premature venting of 
gases with a 32-in stemming region. It 
is concluded then that for the conditions 
of these tests, a stemming length- 
to-charge-diameter ratio of 26 (32 in of 
stemming) was adequate to prevent igni- 
tion of methane. However, a stemming 
length of 16 charge diameters (20 in of 
stemming) could have resulted in methane 
ignitions because of early venting of hot 
gases through fractures in the limestone 
rock. 

Bauer, Calder & Workman, Inc. , sug- 
gested a simple physical model to predict 
the time required to eject stemming. 
This model depends only on inertia of the 
stemming material and not on frictional 
forces to resist movement, and thus the 
acceleration, a, of the stemming is given 
by 

a=|, (4) 

where F is the force exerted on the stem- 
ming by explosive gases and M is the mass 
of the stemming. 
The equation of motion is thus 



S = V t + 1/2 a t 



(5) 



where S is distance traveled by the stem- 
ming, V is initial velocity, and t 
= time in seconds. 
Combining equations with V = gives 



FIGURE 5.— Observed burden velocities versus length of stem- 
ming for the 1 Vi-in-diameter shot series. 



-V¥ 



(6) 



113 



The force F can be estimated from the 
borehole pressure, P, times the cross- 
sectional area of the hole, A, or 



F = PA, 
and the mass of stemming equals 
M = p s M, 



(7) 



(8) 



where Ps is the density of stemming ma- 
terial, A is the cross-sectional area, 
and I is the length of stemming. 
Substituting gives 



t = 



2S p s I . 
P 



(9) 



This prediction method yields ejection 
times for the shots in this investiga- 
tion as shown in table 3. Also shown in 
table 3 is the time of first observed 
stemming movement and observed time to 
ejection. 



All observed times for ejection were 
much longer than the calculated times. 
In many cases, no stemming movement was 
observed before the calculated time was 
past. This calculation method then 
should be used only to obtain an estimate 
of the minimum stemming ejection time. 
Improved estimates will require the 
addition of frictional forces in the 
calculations. 

A clue to this late stemming movement 
is provided by the instrumentation used 
in a 6-in vertical cratering blast by 
Sandia National Laboratories 4 (shot V-l). 
The blasthole was instrumented with the 
SLIFER system to monitor the detonation 
rate of the explosive. This instrument 

4 Shirey, D. L., J. E. Uhl, and R. L. 
Parrish. Atlas Cratering Tests. Ch. in 
Oil Shale Program Quarterly Reports, Oct- 
ober 1984 through March 1985. Sandia 
Nat. Lab., Rep. SAND 85-2768, 1986, 
pp. 5-26. 




FIGURE 6.— SLIFER data from hole 3 of shot V-1 showing detonation of explosive column and crushing rate of stemming. From 
Sandia National Laboratories. 



114 



TABLE 3. - Calculated and actual stemming ejection times 





Hole 


Stemming 


Calculated 


First stemming 


Observed 


Shot 


diameter, in 


length, in 


ejection 
t ime , ms 


movement , ms 


ejection 
time, ras 


S-3 


1-1/2 


20 


0.5 


3.4 


13 


S-4 


1-1/2 


20 


.5 


4.6 


9 


S-6 


1-1/2 


20 


.6 


6.1 


32 


V-1 1 


6 


108 


2.1 


10 


Retained 



6-in cratering shot from Sandia National Laboratories. 



also observed the rate of crushing or 
compaction of the stemming in the bore- 
hole. The record for shot V-1 is shown 
in figure 6. The first 6 ft of the rec- 
ord shows the detonation of the explo- 
sive. The detonation velocity is approx- 
imately 19,000 ft/s. Above 6 ft, the 
record shows the crushing of the stem- 
ming. This crushing extends to 12-1/2 
ft, or 2-1/2 ft from the hole collar, and 



takes 7.5 ms to complete. The time taken 
for the crushing to propagate through the 
stemming is close to but somewhat less 
than the observed time for the start of 
stemming movement. The stemming appears 
to be bridging in the hole and preventing 
movement until it is crushed by the pres- 
sure pulse in the stemming column, and 
only then does it start to move. 



CONCLUSIONS 



High-speed films of single-hole crater 
test blasts in two surface limestone 
quarries were analyzed to evaluate the 
ability of stemming to contain explosive 
gases. Stemming ejection and burden mo- 
tions were examined. When sufficient 
stemming was used, ejection of stemming 
was prevented. A ratio of length of 
stemming to charge diameter of 26 or more 
was found to prevent premature ejection 
of stemming and venting of gases. 

If release of stemming does occur, the 
time required for stemming ejection may 
be sufficient to permit burden movement 
to start, with expansion of explosive 
gases into the fractured burden and cool- 
ing of the explosive gases. An estimate 
of this cooling was made using thermo- 
dynamic principles. In three tests with 
a ratio of stemming length to charge di- 
ameter of 16, the explosive gases cooled 
below the ignition temperture of methane 
in the time required for stemming 
ejection, but venting of gases through 



fractures occurred before stemming ejec- 
tion and the temperature of the vented 
gas was estimated to be above the methane 
ignition temperature. For the conditions 
of these tests, it is thus concluded that 
a stemming length of 16 charge diameters 
could have resulted in methane ignition. 

Stemming ejection takes much longer and 
is more complex than would be predicted 
by a simple calculation based on inertia 
of the stemming material. Ejection times 
were three times or more longer than 
those predicted by a simple inertia 
model. Reasons for this are the addi- 
tional time required for the stress wave 
to crush the stemming material and cause 
it to start to move and the subsequent 
decrease of borehole pressure through 
crushing and expansion of the bore- 
hole. A better understanding of these 
mechanisms requires further research with 
more sophisticated instrumentation, such 
as the SLIFER system. 



U S. GOVERNMENT PRINTING OFFICE 1987 605-01760024 



INT.-BU.0F MINES,PGH.,PA. 28478 



C 214 



— 



^— 



a* .VSSfcy. ^ a> ♦: 






• o' 













^ V % 













>-^ 












■s. O 













> O* V M 













V» 


























3^, 



^ < f 



^v 1 



^0< 









\* 









*' ** 
















m- "♦*<? : 




A 1, ..*'•« <** 






A v «i 



^^' 















A*"%. "^^^^ c^^ : ^S^ r » a>*^ 



^0^ 



^°-v 







« .4S 



.** 



"oV 










^0* 




* K 1 

- V''^.^-^ "•■• v ..... V" 



V*^V v^v v^v 







*° A* 






^0* •« 



tf°- 




©♦. ?••* *P V *-vT« <# ©,. *5vV a©' 



J?**, 



0* .M** "*b a^ .«•'■ 



Ati&X. .*\>$&S /s«^%, ^ 







'bV 



^- ^o< :-^IS; «b* :SH%£- "*+#' i^M:. "^ ?&&*:+ ^^ •% 



«5^ 



V /°^ v. 






%^ 



/.&> /\^^\ ^^%>o /.C^\ / 




• ^0 











* A.V "5-. s H7ffl\ff* a5> 




• * * ° - Ox 



*► ^ 



^ 



^ ** 




BOOKBINDING I .^ ^x * • . - * aO ^ * • « ' V" ©♦ *.,i' a 



v »v 



*^ 



• *^ "» «VdJJV * a •a'* *6 A *^mO»«^' v ^ ** * •* 









LIBRARY OF CONGRESS 



002 951 005 2 



